Search Results

Search found 96005 results on 3841 pages for 'user group'.

Page 524/3841 | < Previous Page | 520 521 522 523 524 525 526 527 528 529 530 531  | Next Page >

  • RDC stops working after a period of time.

    - by xjerx
    I have a workstation with RDC configured for the employee. When they leave at the end of their day they lock the pc (windows key + l). They go home connect to our VPN and log back in. Everything works fine. The following morning they will attempt to log in before they return to the office. The computer does not respond to the RDC request. I've found that it becomes completely inactive to any ICMP requests. Once the user reboots the computer everything works fine again. I'm going to turn off RDC, reboot, turn RDC back on and reboot again to see if it fixes the problem. Until then does anyone have any other ideas?

    Read the article

  • Why do 2 excel (2003) files that are exactly the same have different file size?

    - by meme
    I have two excel files that are exactly the same (in terms of the content of the file) but differ by quite a margin on filesize. One file is 37.5Kb while the other is 56Kb. The only difference I can see is the filename's. I don't know why there is such a big difference. Is there some sort of history or something that is stored with the file that is not visible to the user? If so, how would you delete this? Thanks for your help.

    Read the article

  • Add/Remove script for local printer

    - by GxFlint
    I have a Windows XP machine that runs two applications and both print on a thermal printer connected by a serial port. For one application, the "Generic / Text Only" printer must be present, for the other to work I need to remove it. I've found a few .vbs scripts, but they are for network printer. How do I make them work with my local printer? Is there a better solution? The user would have to run the script every time he needs to switch from an application to another.

    Read the article

  • Cannot Display Chinese Character in my PHP code

    - by Jun1st
    I want to display my Twitter Info in my blog. So I write some code to get it. the issue I got is that Chinese characters displayed as unknown code. Here is the test code. Could anyone take a look and help? Thanks <html> <title>Twitter Test</title> <body> <?php function mystique_objectToArray($object){ if(!is_object($object) && !is_array($object)) return $object; if(is_object($object)) $object = get_object_vars($object); return array_map('mystique_objectToArray', $object); } define( 'ABSPATH', dirname(dirname(__FILE__)) . '/' ); require_once('/home/jun1st/jun1stfeng.com/wp-includes/class-snoopy.php'); $snoopy = new Snoopy; $response = @$snoopy->fetch("http://twitter.com/users/show/jun1st.json"); if ($response) $userdata = json_decode($snoopy->results, true); else $error = true; $response = @$snoopy->fetch("http://twitter.com/statuses/user_timeline/jun1st.json"); if ($response) $tweets = json_decode($snoopy->results, true); else $error = true; if(!$error): // for php < 5 (included JSON returns object) $userdata = mystique_objectToArray($userdata); $tweets = mystique_objectToArray($tweets); $twitdata = array(); $twitdata['user']['profile_image_url'] = $userdata['profile_image_url']; $twitdata['user']['name'] = $userdata['name']; $twitdata['user']['screen_name'] = $userdata['screen_name']; $twitdata['user']['followers_count'] = $userdata['followers_count']; $i = 0; foreach($tweets as $tweet): $twitdata['tweets'][$i]['text'] = $tweet['text']; $twitdata['tweets'][$i]['created_at'] = $tweet['created_at']; $twitdata['tweets'][$i]['id'] = $tweet['id']; $i++; endforeach; endif; // only show if the twitter data from the database is newer than 6 hours if(is_array($twitdata['tweets'])): ?> <div class="clear-block"> <div class="avatar"><img src="<?php echo $twitdata['user']['profile_image_url']; ?>" alt="<?php echo $twitdata['user']['name']; ?>" /></div> <div class="info"><a href="http://www.twitter.com/jun1st"><?php echo $twitdata['user']['name']; ?> </a><br /><span class="followers"> <?php printf(__("%s followers","mystique"),$twitdata['user']['followers_count']); ?></span></div> </div> <ul> <?php $i = 0; foreach($twitdata['tweets'] as $tweet): $pattern = '/\@(\w+)/'; $replace = '<a rel="nofollow" href="http://twitter.com/$1">@$1</a>'; $tweet['text'] = preg_replace($pattern, $replace , $tweet['text']); $tweet['text'] = make_clickable($tweet['text']); // remove +XXXX $tweettime = substr_replace($tweet['created_at'],'',strpos($tweet['created_at'],"+"),5); $link = "http://twitter.com/".$twitdata['user']['screen_name']."/statuses/".$tweet['id']; echo '<li><span class="entry">' . $tweet['text'] .'<a class="date" href="'.$link.'" rel="nofollow">'.$tweettime.'</a></span></li>'; $i++; if ($i == $twitcount) break; endforeach; ?> </ul> <? endif?> ?> </body> </html>

    Read the article

  • Eager/Lazy loaded member always empty with JPA one-to-many relationship

    - by Kaleb Pederson
    I have two entities, a User and Role with a one-to-many relationship from user to role. Here's what the tables look like: mysql> select * from User; +----+-------+----------+ | id | name | password | +----+-------+----------+ | 1 | admin | admin | +----+-------+----------+ 1 row in set (0.00 sec) mysql> select * from Role; +----+----------------------+---------------+----------------+ | id | description | name | summary | +----+----------------------+---------------+----------------+ | 1 | administrator's role | administrator | Administration | | 2 | editor's role | editor | Editing | +----+----------------------+---------------+----------------+ 2 rows in set (0.00 sec) And here's the join table that was created: mysql> select * from User_Role; +---------+----------+ | User_id | roles_id | +---------+----------+ | 1 | 1 | | 1 | 2 | +---------+----------+ 2 rows in set (0.00 sec) And here's the subset of orm.xml that defines the tables and relationships: <entity class="User" name="User"> <table name="User" /> <attributes> <id name="id"> <generated-value strategy="AUTO" /> </id> <basic name="name"> <column name="name" length="100" unique="true" nullable="false"/> </basic> <basic name="password"> <column length="255" nullable="false" /> </basic> <one-to-many name="roles" fetch="EAGER" target-entity="Role" /> </attributes> </entity> <entity class="Role" name="Role"> <table name="Role" /> <attributes> <id name="id"> <generated-value strategy="AUTO"/> </id> <basic name="name"> <column name="name" length="40" unique="true" nullable="false"/> </basic> <basic name="summary"> <column name="summary" length="100" nullable="false"/> </basic> <basic name="description"> <column name="description" length="255"/> </basic> </attributes> </entity> Yet, despite that, when I retrieve the admin user, I get back an empty collection. I'm using Hibernate as my JPA provider and it shows the following debug SQL: select user0_.id as id8_, user0_.name as name8_, user0_.password as password8_ from User user0_ where user0_.name=? limit ? When the one-to-many mapping is lazy loaded, that's the only query that's made. This correctly retrieves the one admin user. I changed the relationship to use eager loading and then the following query is made in addition to the above: select roles0_.User_id as User1_1_, roles0_.roles_id as roles2_1_, role1_.id as id9_0_, role1_.description as descript2_9_0_, role1_.name as name9_0_, role1_.summary as summary9_0_ from User_Role roles0_ left outer join Role role1_ on roles0_.roles_id=role1_.id where roles0_.User_id=? Which results in the following results: +----------+-----------+--------+----------------------+---------------+----------------+ | User1_1_ | roles2_1_ | id9_0_ | descript2_9_0_ | name9_0_ | summary9_0_ | +----------+-----------+--------+----------------------+---------------+----------------+ | 1 | 1 | 1 | administrator's role | administrator | Administration | | 1 | 2 | 2 | editor's role | editor | Editing | +----------+-----------+--------+----------------------+---------------+----------------+ 2 rows in set (0.00 sec) Hibernate obviously knows about the roles, yet getRoles() still returns an empty collection. Hibernate also recognized the relationship sufficiently to put the data in the first place. What problems can cause these symptoms?

    Read the article

  • Neo4j increasing latency as SKIP increases on Cypher query + REST API

    - by voldomazta
    My setup: Java(TM) SE Runtime Environment (build 1.7.0_45-b18) Java HotSpot(TM) 64-Bit Server VM (build 24.45-b08, mixed mode) Neo4j 2.0.0-M06 Enterprise First I made sure I warmed up the cache by executing the following: START n=node(*) RETURN COUNT(n); START r=relationship(*) RETURN count(r); The size of the table is 63,677 nodes and 7,169,995 relationships Now I have the following query: START u1=node:node_auto_index('uid:39') MATCH (u1:user)-[w:WANTS]->(c:card)<-[h:HAS]-(u2:user) WHERE u2.uid <> 39 WITH u2.uid AS uid, (CASE WHEN w.qty < h.qty THEN w.qty ELSE h.qty END) AS have RETURN uid, SUM(have) AS total ORDER BY total DESC SKIP 0 LIMIT 25 This UID has about 40k+ results that I want to be able to put a pagination to. The initial skip was around 773ms. I tried page 2 (skip 25) and the latency was around the same even up to page 500 it only rose up to 900ms so I didn't really bother. Now I tried some fast forward paging and jumped by thousands so I did 1000, then 2000, then 3000. I was hoping the ORDER BY arrangement will already have been cached by Neo4j and using SKIP will just move to that index in the result and wont have to iterate through each one again. But for each thousand skip I made the latency increased by alot. It's not just cache warming because for one I already warmed up the cache and two, I tried the same skip a couple of times for each skip and it yielded the same results: SKIP 0: 773ms SKIP 1000: 1369ms SKIP 2000: 2491ms SKIP 3000: 3899ms SKIP 4000: 5686ms SKIP 5000: 7424ms Now who the hell would want to view 5000 pages of results? 40k even?! :) Good point! I will probably put a cap on the maximum results a user can view but I was just curious about this phenomenon. Will somebody please explain why Neo4j seems to be re-iterating through stuff which appears to be already known to it? Here is my profiling for the 0 skip: ==> ColumnFilter(symKeys=["uid", " INTERNAL_AGGREGATE65c4d6a2-1930-4f32-8fd9-5e4399ce6f14"], returnItemNames=["uid", "total"], _rows=25, _db_hits=0) ==> Slice(skip="Literal(0)", _rows=25, _db_hits=0) ==> Top(orderBy=["SortItem(Cached( INTERNAL_AGGREGATE65c4d6a2-1930-4f32-8fd9-5e4399ce6f14 of type Any),false)"], limit="Add(Literal(0),Literal(25))", _rows=25, _db_hits=0) ==> EagerAggregation(keys=["uid"], aggregates=["( INTERNAL_AGGREGATE65c4d6a2-1930-4f32-8fd9-5e4399ce6f14,Sum(have))"], _rows=41659, _db_hits=0) ==> ColumnFilter(symKeys=["have", "u1", "uid", "c", "h", "w", "u2"], returnItemNames=["uid", "have"], _rows=146826, _db_hits=0) ==> Extract(symKeys=["u1", "c", "h", "w", "u2"], exprKeys=["uid", "have"], _rows=146826, _db_hits=587304) ==> Filter(pred="((NOT(Product(u2,uid(0),true) == Literal(39)) AND hasLabel(u1:user(0))) AND hasLabel(u2:user(0)))", _rows=146826, _db_hits=146826) ==> TraversalMatcher(trail="(u1)-[w:WANTS WHERE (hasLabel(NodeIdentifier():card(1)) AND hasLabel(NodeIdentifier():card(1))) AND true]->(c)<-[h:HAS WHERE (NOT(Product(NodeIdentifier(),uid(0),true) == Literal(39)) AND hasLabel(NodeIdentifier():user(0))) AND true]-(u2)", _rows=146826, _db_hits=293696) And for the 5000 skip: ==> ColumnFilter(symKeys=["uid", " INTERNAL_AGGREGATE99329ea5-03cd-4d53-a6bc-3ad554b47872"], returnItemNames=["uid", "total"], _rows=25, _db_hits=0) ==> Slice(skip="Literal(5000)", _rows=25, _db_hits=0) ==> Top(orderBy=["SortItem(Cached( INTERNAL_AGGREGATE99329ea5-03cd-4d53-a6bc-3ad554b47872 of type Any),false)"], limit="Add(Literal(5000),Literal(25))", _rows=5025, _db_hits=0) ==> EagerAggregation(keys=["uid"], aggregates=["( INTERNAL_AGGREGATE99329ea5-03cd-4d53-a6bc-3ad554b47872,Sum(have))"], _rows=41659, _db_hits=0) ==> ColumnFilter(symKeys=["have", "u1", "uid", "c", "h", "w", "u2"], returnItemNames=["uid", "have"], _rows=146826, _db_hits=0) ==> Extract(symKeys=["u1", "c", "h", "w", "u2"], exprKeys=["uid", "have"], _rows=146826, _db_hits=587304) ==> Filter(pred="((NOT(Product(u2,uid(0),true) == Literal(39)) AND hasLabel(u1:user(0))) AND hasLabel(u2:user(0)))", _rows=146826, _db_hits=146826) ==> TraversalMatcher(trail="(u1)-[w:WANTS WHERE (hasLabel(NodeIdentifier():card(1)) AND hasLabel(NodeIdentifier():card(1))) AND true]->(c)<-[h:HAS WHERE (NOT(Product(NodeIdentifier(),uid(0),true) == Literal(39)) AND hasLabel(NodeIdentifier():user(0))) AND true]-(u2)", _rows=146826, _db_hits=293696) The only difference is the LIMIT clause on the Top function. I hope we can make this work as intended, I really don't want to delve into doing an embedded Neo4j + my own Jetty REST API for the web app.

    Read the article

  • How to handle recurring dates (dates only) in .NET?

    - by Wayne M
    I am trying to figure out a good way to handle recurring events in .NET, specifically for an ASP.NET MVC application. The idea is that a user can create an event and specify that the event can occur repeatedly after a specific interval (e.g. "every two weeks", "once a month" and so on). What would be the best way to tackle this? My brainstorming right now is to have two tables: Job and RecurringJob. Job is the "master" record and has the description of the job as well a key to what customer it's for, while RecurringJob links back to Job and has additional info on what the occurrence frequency is (e.g. 1 for "once a month") as well as the timespan (e.g. "Weekly", "Monthly"). The issue is how to determine and set the next occurrence of the job since this will have to be something that's done regularly. I've seen two trains of thought with this: This logic should either be stored in a database column and periodically updated, or calculated on the fly in the code. Any thoughts or suggestions on tackling this? Edit: this is for a subscription based web app I'm creating to let service businesses schedule their common recurring jobs easily and track their customers. So a typical use might be to create a "Cut lawn" job for Mr Smith that occurs every month The exact date isn't important - it's the ability for the customer to see that Mr Smith gets his lawn cut every month and followup with him about it. Let me rephrase the above to better convey my idea. A sample use case for the application might be as follows: User pulls up the customer record for John Smith and clicks the Add Job link. The user fills out the form to create a job with a name of "Cut lawn", a start date of 11/15/2009, and selects a checkbox indicating that this job continually occurs. The user is presented with a secondary screen asking for the job frequency. The user indicates (haven't decided how at this point - let's assume select lists) that the job occurs once a month. User clicks save. Now, when the user views the record for John Smith, they can see that he has a job, "Cut lawn", that occurs every month starting from 11/15/2009. On the main dashboard when it's one week prior to the assumed start date, the user sees the job displayed with an indicator such as "12/15/2009 - Cut lawn (John Smith)". A week before the due date someone from the company calls him up to schedule and he says he's going to be out of town until 1/1/2010, so he wants his appointment rescheduled for that date. Our user can change the date for the job to be 1/1/2010, and now the recurrence will start one month from that date (e.g. next time will be 2/1/2010). The idea behind this is that the app is targeting businesses like lawn care, plumbers, carpet cleaners and the like where the exact date isn't as important (because it can and will change as people are busy), the key thing is to give the business an indicator that Mr. Smith's monthly service is coming up, and someone should give him a call to determine when exactly it can be scheduled for. In effect give these businesses a way to track repeat business and know when it's time to followup with a customer.

    Read the article

  • onCommand on button not firing inside gridView??

    - by sah302
    This is really driving me crazy. I've got a button inside a gridview to remove that item from the gridview (its datasource is a list). I've got the list being saved to session anytime a change is being made to it, and on page_load check if that session variable is empty, if not, then set that list to bind to the gridview. Code Behind: Public accomplishmentTypeDao As New AccomplishmentTypeDao() Public accomplishmentDao As New AccomplishmentDao() Public userDao As New UserDao() Public facultyDictionary As New Dictionary(Of Guid, String) Public facultyList As New List(Of User) Public associatedFaculty As New List(Of User) Public facultyId As New Guid Protected Sub Page_Load(ByVal sender As Object, ByVal e As System.EventArgs) Handles Me.Load 'If Not Session("associatedFaculty") Is Nothing Then' ' Dim associatedFacultyArray As User() = DirectCast(Session("associatedFaculty"), User())' ' associatedFaculty = associatedFacultyArray.ToList()' 'End If' Page.Title = "Add a New Faculty Accomplishment" ddlAccomplishmentType.DataSource = accomplishmentTypeDao.getEntireTable() ddlAccomplishmentType.DataTextField = "Name" ddlAccomplishmentType.DataValueField = "Id" ddlAccomplishmentType.DataBind() facultyList = userDao.getListOfUsersByUserGroupName("Faculty") For Each faculty As User In facultyList facultyDictionary.Add(faculty.Id, faculty.LastName & ", " & faculty.FirstName) Next If Not Page.IsPostBack Then ddlFacultyList.DataSource = facultyDictionary ddlFacultyList.DataTextField = "Value" ddlFacultyList.DataValueField = "Key" ddlFacultyList.DataBind() End If gvAssociatedUsers.DataSource = associatedFaculty gvAssociatedUsers.DataBind() End Sub Protected Sub deleteUser(ByVal sender As Object, ByVal e As System.Web.UI.WebControls.CommandEventArgs) facultyId = New Guid(e.CommandArgument.ToString()) associatedFaculty.Remove(associatedFaculty.Find(Function(user) user.Id = facultyId)) Session("associatedFaculty") = associatedFaculty.ToArray() gvAssociatedUsers.DataBind() upAssociatedFaculty.Update() End Sub Protected Sub btnAddUser_Click(ByVal sender As Object, ByVal e As EventArgs) Handles btnAddUser.Click facultyId = New Guid(ddlFacultyList.SelectedValue) associatedFaculty.Add(facultyList.Find(Function(user) user.Id = facultyId)) Session.Add("associatedFaculty", associatedFaculty.ToArray()) gvAssociatedUsers.DataBind() upAssociatedFaculty.Update() End Sub Protected Sub Delete(ByVal sender As Object, ByVal e As System.Web.UI.WebControls.CommandEventArgs) End Sub End Class Markup: <asp:UpdatePanel ID="upAssociatedFaculty" runat="server" UpdateMode="Conditional"> <ContentTemplate> <p><b>Created By:</b> <asp:Label ID="lblCreatedBy" runat="server"></asp:Label></p> <p><b>Accomplishment Type: </b><asp:DropDownList ID="ddlAccomplishmentType" runat="server"></asp:DropDownList></p> <p><b>Accomplishment Applies To: </b><asp:DropDownList ID="ddlFacultyList" runat="server"></asp:DropDownList> &nbsp;<asp:Button ID="btnAddUser" runat="server" Text="Add Faculty" OnClientClick="incrementCounter();" /></p> <p> <asp:GridView ID="gvAssociatedUsers" runat="server" AutoGenerateColumns="false" GridLines="None" ShowHeader="false"> <Columns> <asp:BoundField DataField="Id" HeaderText="Id" Visible="False" /> <asp:TemplateField ShowHeader="False"> <ItemTemplate> <span style="margin-left: 15px;"> <p><%#Eval("LastName")%>, <%#Eval("FirstName")%> <asp:Button ID="btnUnassignUser" runat="server" CausesValidation="false" CommandArgument='<%# Eval("Id") %>' CommandName="Delete" OnCommand="deleteUser" Text='Remove' /></p> </span> </ItemTemplate> </asp:TemplateField> </Columns> <EmptyDataTemplate> <em>There are currently no faculty associated with this accomplishment.</em> </EmptyDataTemplate> </asp:GridView> </p> </ContentTemplate> </asp:UpdatePanel> Now here is the crazy part I am simply boggled by, if I uncomment the If Not Session... block of page_load, then deleteUser will never fire when btnUnassignUser is clicked. If I keep it commented out...it fires no problem, but then of course my list can never have more than one item since I am not loading the saved list from session into the gridview but just a fresh one. But the button click is being registered, because page_load is being stepped through again when I am viewing in debug mode, just deleteUser never fires. Why is this happening?? And how can I fix it??

    Read the article

  • method is not called from xhtml

    - by Amlan Karmakar
    Whenever I am clicking the h:commandButton,the method associated with the action is not called.action="${statusBean.update}" is not working, the update is not being called. 1) Here is my xhtml page <!DOCTYPE html PUBLIC "-//W3C//DTD XHTML 1.0 Transitional//EN" "http://www.w3.org/TR/xhtml1/DTD/xhtml1-transitional.dtd"> <html xmlns="http://www.w3.org/1999/xhtml" xmlns:ui="http://java.sun.com/jsf/facelets" xmlns:h="http://java.sun.com/jsf/html" xmlns:f="http://java.sun.com/jsf/core" xmlns:p="http://primefaces.org/ui"> <h:head></h:head> <h:body> <h:form > <p:dataList value="#{statusBean.statusList}" var="p"> <h:outputText value="#{p.statusId}-#{p.statusmsg}"/><br/> <p:inputText value="#{statusBean.comment.comment}"/> <h:commandButton value="comment" action="${statusBean.update}"></h:commandButton> </p:dataList> </h:form> </h:body> </html> 2)Here is my statusBean package com.bean; import java.util.List; import javax.faces.context.FacesContext; import javax.persistence.EntityManager; import javax.persistence.EntityManagerFactory; import javax.persistence.Persistence; import javax.persistence.Query; import javax.servlet.http.HttpSession; import com.entity.Album; import com.entity.Comment; import com.entity.Status; import com.entity.User; public class StatusBean { Comment comment; Status status; private EntityManager em; public Comment getComment() { return comment; } public void setComment(Comment comment) { this.comment = comment; } public Status getStatus() { return status; } public void setStatus(Status status) { this.status = status; } public StatusBean(){ comment = new Comment(); status=new Status(); EntityManagerFactory emf=Persistence.createEntityManagerFactory("FreeBird"); em =emf.createEntityManager(); } public String save(){ FacesContext context = FacesContext.getCurrentInstance(); HttpSession session = (HttpSession) context.getExternalContext().getSession(true); User user = (User) session.getAttribute("userdet"); status.setEmail(user.getEmail()); System.out.println("status save called"); em.getTransaction().begin(); em.persist(status); em.getTransaction().commit(); return "success"; } public List<Status> getStatusList(){ FacesContext context = FacesContext.getCurrentInstance(); HttpSession session = (HttpSession) context.getExternalContext().getSession(true); User user=(User) session.getAttribute("userdet"); Query query = em.createQuery("SELECT s FROM Status s WHERE s.email='"+user.getEmail()+"'", Status.class); List<Status> results =query.getResultList(); return results; } public String update(){ System.out.println("Update Called..."); //comment.setStatusId(Integer.parseInt(statusId)); em.getTransaction().begin(); em.persist(comment); em.getTransaction().commit(); return "success"; } }

    Read the article

  • Clearcase - selective merge.

    - by Keshav
    Hi, I have a peculiar Clearcase doubt. I cannot fully describe why I'm doing such a confusing architecture, but I need to do it (thanks to the mistake done by someone long back). Ok, here's a bit of detail: B1 is a contaminated branch where both my group's changes and another group's changes got mixed together so badly that there is no way of finding which code is whose). So the solution proposed is to create a new branch called B2 (at the same level as B1) and put all the unmodified code of the other group on it (The way to do that would be to merge B1 with B2 and then go about removing all changes from it till it becomes original). Then create a CR branch on B1 and keep only my group's newly added files or modified files on that branch. Finally create an integration branch out of B2 and merge the changes from CR branch of B1 to integration branch of B2. So here is what I did: (The use case is where I have dir D where file a, b and c are there. My group ended up modifying file a while b and c are not modified at all). There is a branch B1 on which there are files a, b and c. There is another branch B2. A merge is done from B1 to B2. Now B2 also has a, b and c. At this point both branch B1 and B2 are same. Now I delete file a from branch B2 (rmname). Now B2 has b and c only. I put a label to this branch called Label1. This makes the code with label Label1 as the unmodified code from other group. Now I create a sub branch called CR1 from B1 and delete all the files that are there in B2 branch (i.e b and c) such that it contains only the modified code from original code on it. In my case it is file a. At this point branch B2 with label Label1 has files b and c (those are unmodified code) and branch CR1 coming off B1 has only a (that is modified by us). Now I create another branch called integration branch that comes off B2 Label1. And then I do a merge of CR branch on to that expecting that it will have all three files a, b and c for me. All I'd need to do is to do a version tree view and see who modified what. But the problem I face is that since I had done a rmname of file a on branch B2 earlier to putting Label. The merge does not really take the file a from CR branch. How to I get around that problem. I want to selectively merge. Is it possible? sorry if it is a bad design. I'm not really conversant with Clear case and have limited options and time to clear some one else's mess.

    Read the article

  • Do I must expose the aggregate children as public properties to implement the Persistence ignorance?

    - by xuehua
    Hi all, I'm very glad that i found this website recently, I've learned a lot from here. I'm from China, and my English is not so good. But i will try to express myself what i want to say. Recently, I've started learning about Domain Driven Design, and I'm very interested about it. And I plan to develop a Forum website using DDD. After reading lots of threads from here, I understood that persistence ignorance is a good practice. Currently, I have two questions about what I'm thinking for a long time. Should the domain object interact with repository to get/save data? If the domain object doesn't use repository, then how does the Infrastructure layer (like unit of work) know which domain object is new/modified/removed? For the second question. There's an example code: Suppose i have a user class: public class User { public Guid Id { get; set; } public string UserName { get; set; } public string NickName { get; set; } /// <summary> /// A Roles collection which represents the current user's owned roles. /// But here i don't want to use the public property to expose it. /// Instead, i use the below methods to implement. /// </summary> //public IList<Role> Roles { get; set; } private List<Role> roles = new List<Role>(); public IList<Role> GetRoles() { return roles; } public void AddRole(Role role) { roles.Add(role); } public void RemoveRole(Role role) { roles.Remove(role); } } Based on the above User class, suppose i get an user from the IUserRepository, and add an Role for it. IUserRepository userRepository; User user = userRepository.Get(Guid.NewGuid()); user.AddRole(new Role() { Name = "Administrator" }); In this case, i don't know how does the repository or unit of work can know that user has a new role? I think, a real persistence ignorance ORM framework should support POCO, and any changes occurs on the POCO itself, the persistence framework should know automatically. Even if change the object status through the method(AddRole, RemoveRole) like the above example. I know a lot of ORM can automatically persistent the changes if i use the Roles property, but sometimes i don't like this way because of the performance reason. Could anyone give me some ideas for this? Thanks. This is my first question on this site. I hope my English can be understood. Any answers will be very appreciated.

    Read the article

  • Real Excel Templates I

    - by Tim Dexter
    As promised, I'm starting to document the new Excel templates that I teased you all with a few weeks back. Leslie is buried in 11g documentation and will not get to officially documenting the templates for a while. I'll do my best to be professional and not ramble on about this and that, although the weather here has finally turned and its 'scorchio' here in Colorado today. Maybe our stand of Aspen will finally come into leaf ... but I digress. Preamble These templates are not actually that new, I helped in a small way to develop them a few years back with Excel 'meistress' Shirley for a company that was trying to use the Report Manager(RR) Excel FSG outputs under EBS 12. The functionality they needed was just not there in the RR FSG templates, the templates are actually XSL that is created from the the RR Excel template builder and fed to BIP for processing. Think of Excel from our RTF templates and you'll be there ie not really Excel but HTML masquerading as Excel. Although still under controlled release in EBS they have now made their way to the standlone release and are willing to share their Excel goodness. You get everything you have with hte Excel Analyzer Excel templates plus so much more. Therein lies a question, what will happen to the Analyzer templates? My understanding is that both will come together into a single Excel template format some time in the post-11g release world. The new XLSX format for Exce 2007/10 is also in the mix too so watch this space. What more do these templates offer? Well, you can structure data in the Excel output. Similar to RTF templates you can create sheets of data that have master-detail n relationships. Although the analyzer templates can do this, you have to get into macros whereas BIP will do this all for you. You can also use native XSL functions in your data to manipulate it prior to rendering. BP functions are not currently supported. The most impressive, for me at least, is the sheet 'bursting'. You can split your hierarchical data across multiple sheets and dynamically name those sheets. Finally, you of course, still get all the native Excel functionality. Pre-reqs You must be on 10.1.3.4.1 plus the latest rollup patch, 9546699. You can patch upa BIP instance running with OBIEE, no problem You need Excel 2000 or above to build the templates Some patience - there is no Excel template builder for these new templates. So its all going to have to be done by hand. Its not that tough but can get a little 'fiddly'. You can not test the template from Excel , it has to be deployed and then run. Limitations The new templates are definitely superior to the Analyzer templates but there are a few limitations. Re-grouping is not supported. You can only follow a data hierarchy not bend it to your will unless you want to get into macros. No support for BIP functions. The templates support native XSL functions only. No template builder Getting Started The templates make the use of named cells and groups of cells to allow BIP to find the insertion point for data points. It also uses a hidden sheet to store calculation mappings from named cells to XML data elements. To start with, in the great BIP tradition, we need some sample XML data. Becasue I wanted to show the master-detail output we need some hierarchical data. If you have not yet gotten into the data templates, now is a good time, I wrote a post a while back starting from the simple to more complex. They generate ideal data sets for these templates. Im working with the following data set: <EMPLOYEES> <LIST_G_DEPT> <G_DEPT> <DEPARTMENT_ID>10</DEPARTMENT_ID> <DEPARTMENT_NAME>Administration</DEPARTMENT_NAME> <LIST_G_EMP> <G_EMP> <EMPLOYEE_ID>200</EMPLOYEE_ID> <EMP_NAME>Jennifer Whalen</EMP_NAME> <EMAIL>JWHALEN</EMAIL> <PHONE_NUMBER>515.123.4444</PHONE_NUMBER> <HIRE_DATE>1987-09-17T00:00:00.000-06:00</HIRE_DATE> <SALARY>4400</SALARY> </G_EMP> </LIST_G_EMP> <TOTAL_EMPS>1</TOTAL_EMPS> <TOTAL_SALARY>4400</TOTAL_SALARY> <AVG_SALARY>4400</AVG_SALARY> <MAX_SALARY>4400</MAX_SALARY> <MIN_SALARY>4400</MIN_SALARY> </G_DEPT> ... <LIST_G_DEPT> <EMPLOYEES> Simple enough to follow and bread and butter stuff for an RTF template. Building the Template For an Excel template we need to start by thinking about how we want to render the data. Come up with a sample output in Excel. Its all dummy data, nothing marked up yet with one row of data for each level. I have the department name and then a repeating row for the employees. You can apply Excel formatting to the layout. The total is going to be derived from a data element. We'll get to Excel functions later. Marking Up Cells Next we need to start marking up the cells with custom names to map them to data elements. The cell names need to follow a specific format: For data grouping, XDO_GROUP_?group_name? For data elements, XDO_?element_name? Notice the question mark delimter, the group_name and element_name are case sensitive. The next step is to find how to name cells; the easiest method is to highlight the cell and then type in the name. You can also find the Name Manager dialog. I use 2007 and its available on the ribbon under the Formulas section Go thorugh the process of naming all the cells for the element values you have. Using my data set from above.You should end up with something like this in your 'Name Manager' dialog. You can update any mistakes you might have made through this dialog. Creating Groups In the image above you can see there are a couple of named group cells. To create these its a simple case of highlighting the cells that make up the group and then naming them. For the EMP group, highlight the employee row and then type in the name, XDO_GROUP?G_EMP? Notice the 10,000 total is outside of the G_EMP group. Its actually named, XDO_?TOTAL_SALARY?, a query calculated value. For the department group, we need to include the department name cell and the sub EMP grouping and name it, XDO_GROUP?G_DEPT? Notice, the 10,000 total is included in the G_DEPT group. This will ensure it repeats at the department level. Lastly, we do need to include a special sheet in the workbook. We will not have anything meaningful in there for now, but it needs to be present. Create a new sheet and name it XDO_METADATA. The name is important as the BIP rendering engine will looking for it. For our current example we do not need anything other than the required stuff in our XDO_METADATA sheet but, it must be present. Easy enough to hide it. Here's what I have: The only cell that is important is the 'Data Constraints:' cell. The rest is optional. To save curious users getting distracted, hide the metadata sheet. Deploying & Running Templates We should now have a usable Excel template. Loading it into a report is easy enough using the browser UI, just like an RTF template. Set the template type to Excel. You will now be able to run the report and hopefully get something like this. You will not get the red highlighting, thats just some conditional formatting I added to the template using Excel functionality. Your dates are probably going to look raw too. I got around this for now using an Excel function on the cell: =--REPLACE(SUBSTITUTE(E8,"T"," "),LEN(E8)-6,6,"") Google to the rescue on that one. Try some other stuff out. To avoid constantly loading the template through the UI. If you have BIP running locally or you can access the reports repository, once you have loaded the template the first time. Just save the template directly into the report folder. I have put together a sample report using a sample data set, available here. Just drop the xml data file, EmpbyDeptExcelData.xml into 'demo files' folder and you should be good to go. Thats the basics, next we'll start using some XSL functions in the template and move onto the 'bursting' across sheets.

    Read the article

  • Modifying AD Schema permissions from the command line

    - by Ryan Roussel
    Recently while making some changes for a client, I accidently dug myself into a pretty deep hole.  I was trying to explicitly deny a certain user from reading a few group policies including the Default Domain Policy.  When I went in to make the change I accidently denied Authenticated Users rather than the AD user object.  This of course made the GPO inaccessible to all users including any with domain admin rights.  The policy could no longer be modified in the GPMC and worse, changes could not be made through ADSIedit.   The errors I was getting from inside ADSIedit when trying to edit the container looked like this This object has one or more property sheets currently open. Invalid path to object The only solution was to strip Authenticated Users from the container ACL completely in the schema, then re-add it back with the default read and apply rights.  To perform this action, I used a command I had never used before:  DSALCS.exe  It’s part of the DSMOD group of tools.  Since this command interacts with the actual schema, you have to know the full LDAP container or object name.  In this case the GUID of the Default Domain Policy: {31B2F340-016D-11D2-945F-00C04FB984F9}   The actual commands I ran looked like this:   To display the current ACL of the container: c:\>dsacls “cn={31B2F340-016D-11D2-945F-00C04FB984F9},cn=Policies,cn=System, dc=domain,dc=com” /A To strip Authenticated Users from the ACL of the container: c:\>dsacls “cn={31B2F340-016D-11D2-945F-00C04FB984F9},cn=Policies,cn=System, dc=domain,dc=com” /R “NT Authority\Authenticated Users”   For full reference of the DSACLS.EXE command visit: http://support.microsoft.com/kb/281146 Once the Authenticated Users was cleared from the ACL, I was able to use Group Policy Management Console to reassign the default permissions.

    Read the article

  • Entity Association Mapping with Code First Part 1 : Mapping Complex Types

    - by mortezam
    Last week the CTP5 build of the new Entity Framework Code First has been released by data team at Microsoft. Entity Framework Code-First provides a pretty powerful code-centric way to work with the databases. When it comes to associations, it brings ultimate flexibility. I’m a big fan of the EF Code First approach and am planning to explain association mapping with code first in a series of blog posts and this one is dedicated to Complex Types. If you are new to Code First approach, you can find a great walkthrough here. In order to build a solid foundation for our discussion, we will start by learning about some of the core concepts around the relationship mapping.   What is Mapping?Mapping is the act of determining how objects and their relationships are persisted in permanent data storage, in our case, relational databases. What is Relationship mapping?A mapping that describes how to persist a relationship (association, aggregation, or composition) between two or more objects. Types of RelationshipsThere are two categories of object relationships that we need to be concerned with when mapping associations. The first category is based on multiplicity and it includes three types: One-to-one relationships: This is a relationship where the maximums of each of its multiplicities is one. One-to-many relationships: Also known as a many-to-one relationship, this occurs when the maximum of one multiplicity is one and the other is greater than one. Many-to-many relationships: This is a relationship where the maximum of both multiplicities is greater than one. The second category is based on directionality and it contains two types: Uni-directional relationships: when an object knows about the object(s) it is related to but the other object(s) do not know of the original object. To put this in EF terminology, when a navigation property exists only on one of the association ends and not on the both. Bi-directional relationships: When the objects on both end of the relationship know of each other (i.e. a navigation property defined on both ends). How Object Relationships Are Implemented in POCO domain models?When the multiplicity is one (e.g. 0..1 or 1) the relationship is implemented by defining a navigation property that reference the other object (e.g. an Address property on User class). When the multiplicity is many (e.g. 0..*, 1..*) the relationship is implemented via an ICollection of the type of other object. How Relational Database Relationships Are Implemented? Relationships in relational databases are maintained through the use of Foreign Keys. A foreign key is a data attribute(s) that appears in one table and must be the primary key or other candidate key in another table. With a one-to-one relationship the foreign key needs to be implemented by one of the tables. To implement a one-to-many relationship we implement a foreign key from the “one table” to the “many table”. We could also choose to implement a one-to-many relationship via an associative table (aka Join table), effectively making it a many-to-many relationship. Introducing the ModelNow, let's review the model that we are going to use in order to implement Complex Type with Code First. It's a simple object model which consist of two classes: User and Address. Each user could have one billing address. The Address information of a User is modeled as a separate class as you can see in the UML model below: In object-modeling terms, this association is a kind of aggregation—a part-of relationship. Aggregation is a strong form of association; it has some additional semantics with regard to the lifecycle of objects. In this case, we have an even stronger form, composition, where the lifecycle of the part is fully dependent upon the lifecycle of the whole. Fine-grained domain models The motivation behind this design was to achieve Fine-grained domain models. In crude terms, fine-grained means “more classes than tables”. For example, a user may have both a billing address and a home address. In the database, you may have a single User table with the columns BillingStreet, BillingCity, and BillingPostalCode along with HomeStreet, HomeCity, and HomePostalCode. There are good reasons to use this somewhat denormalized relational model (performance, for one). In our object model, we can use the same approach, representing the two addresses as six string-valued properties of the User class. But it’s much better to model this using an Address class, where User has the BillingAddress and HomeAddress properties. This object model achieves improved cohesion and greater code reuse and is more understandable. Complex Types: Splitting a Table Across Multiple Types Back to our model, there is no difference between this composition and other weaker styles of association when it comes to the actual C# implementation. But in the context of ORM, there is a big difference: A composed class is often a candidate Complex Type. But C# has no concept of composition—a class or property can’t be marked as a composition. The only difference is the object identifier: a complex type has no individual identity (i.e. no AddressId defined on Address class) which make sense because when it comes to the database everything is going to be saved into one single table. How to implement a Complex Types with Code First Code First has a concept of Complex Type Discovery that works based on a set of Conventions. The convention is that if Code First discovers a class where a primary key cannot be inferred, and no primary key is registered through Data Annotations or the fluent API, then the type will be automatically registered as a complex type. Complex type detection also requires that the type does not have properties that reference entity types (i.e. all the properties must be scalar types) and is not referenced from a collection property on another type. Here is the implementation: public class User{    public int UserId { get; set; }    public string FirstName { get; set; }    public string LastName { get; set; }    public string Username { get; set; }    public Address Address { get; set; }} public class Address {     public string Street { get; set; }     public string City { get; set; }            public string PostalCode { get; set; }        }public class EntityMappingContext : DbContext {     public DbSet<User> Users { get; set; }        } With code first, this is all of the code we need to write to create a complex type, we do not need to configure any additional database schema mapping information through Data Annotations or the fluent API. Database SchemaThe mapping result for this object model is as follows: Limitations of this mappingThere are two important limitations to classes mapped as Complex Types: Shared references is not possible: The Address Complex Type doesn’t have its own database identity (primary key) and so can’t be referred to by any object other than the containing instance of User (e.g. a Shipping class that also needs to reference the same User Address). No elegant way to represent a null reference There is no elegant way to represent a null reference to an Address. When reading from database, EF Code First always initialize Address object even if values in all mapped columns of the complex type are null. This means that if you store a complex type object with all null property values, EF Code First returns a initialized complex type when the owning entity object is retrieved from the database. SummaryIn this post we learned about fine-grained domain models which complex type is just one example of it. Fine-grained is fully supported by EF Code First and is known as the most important requirement for a rich domain model. Complex type is usually the simplest way to represent one-to-one relationships and because the lifecycle is almost always dependent in such a case, it’s either an aggregation or a composition in UML. In the next posts we will revisit the same domain model and will learn about other ways to map a one-to-one association that does not have the limitations of the complex types. References ADO.NET team blog Mapping Objects to Relational Databases Java Persistence with Hibernate

    Read the article

  • 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.

    Read the article

  • SignalR Auto Disconnect when Page Changed in AngularJS

    - by Shaun
    Originally posted on: http://geekswithblogs.net/shaunxu/archive/2014/05/30/signalr-auto-disconnect-when-page-changed-in-angularjs.aspxIf we are using SignalR, the connection lifecycle was handled by itself very well. For example when we connect to SignalR service from browser through SignalR JavaScript Client the connection will be established. And if we refresh the page, close the tab or browser, or navigate to another URL then the connection will be closed automatically. This information had been well documented here. In a browser, SignalR client code that maintains a SignalR connection runs in the JavaScript context of a web page. That's why the SignalR connection has to end when you navigate from one page to another, and that's why you have multiple connections with multiple connection IDs if you connect from multiple browser windows or tabs. When the user closes a browser window or tab, or navigates to a new page or refreshes the page, the SignalR connection immediately ends because SignalR client code handles that browser event for you and calls the "Stop" method. But unfortunately this behavior doesn't work if we are using SignalR with AngularJS. AngularJS is a single page application (SPA) framework created by Google. It hijacks browser's address change event, based on the route table user defined, launch proper view and controller. Hence in AngularJS we address was changed but the web page still there. All changes of the page content are triggered by Ajax. So there's no page unload and load events. This is the reason why SignalR cannot handle disconnect correctly when works with AngularJS. If we dig into the source code of SignalR JavaScript Client source code we will find something below. It monitors the browser page "unload" and "beforeunload" event and send the "stop" message to server to terminate connection. But in AngularJS page change events were hijacked, so SignalR will not receive them and will not stop the connection. 1: // wire the stop handler for when the user leaves the page 2: _pageWindow.bind("unload", function () { 3: connection.log("Window unloading, stopping the connection."); 4:  5: connection.stop(asyncAbort); 6: }); 7:  8: if (isFirefox11OrGreater) { 9: // Firefox does not fire cross-domain XHRs in the normal unload handler on tab close. 10: // #2400 11: _pageWindow.bind("beforeunload", function () { 12: // If connection.stop() runs runs in beforeunload and fails, it will also fail 13: // in unload unless connection.stop() runs after a timeout. 14: window.setTimeout(function () { 15: connection.stop(asyncAbort); 16: }, 0); 17: }); 18: }   Problem Reproduce In the codes below I created a very simple example to demonstrate this issue. Here is the SignalR server side code. 1: public class GreetingHub : Hub 2: { 3: public override Task OnConnected() 4: { 5: Debug.WriteLine(string.Format("Connected: {0}", Context.ConnectionId)); 6: return base.OnConnected(); 7: } 8:  9: public override Task OnDisconnected() 10: { 11: Debug.WriteLine(string.Format("Disconnected: {0}", Context.ConnectionId)); 12: return base.OnDisconnected(); 13: } 14:  15: public void Hello(string user) 16: { 17: Clients.All.hello(string.Format("Hello, {0}!", user)); 18: } 19: } Below is the configuration code which hosts SignalR hub in an ASP.NET WebAPI project with IIS Express. 1: public class Startup 2: { 3: public void Configuration(IAppBuilder app) 4: { 5: app.Map("/signalr", map => 6: { 7: map.UseCors(CorsOptions.AllowAll); 8: map.RunSignalR(new HubConfiguration() 9: { 10: EnableJavaScriptProxies = false 11: }); 12: }); 13: } 14: } Since we will host AngularJS application in Node.js in another process and port, the SignalR connection will be cross domain. So I need to enable CORS above. In client side I have a Node.js file to host AngularJS application as a web server. You can use any web server you like such as IIS, Apache, etc.. Below is the "index.html" page which contains a navigation bar so that I can change the page/state. As you can see I added jQuery, AngularJS, SignalR JavaScript Client Library as well as my AngularJS entry source file "app.js". 1: <html data-ng-app="demo"> 2: <head> 3: <script type="text/javascript" src="jquery-2.1.0.js"></script> 1:  2: <script type="text/javascript" src="angular.js"> 1: </script> 2: <script type="text/javascript" src="angular-ui-router.js"> 1: </script> 2: <script type="text/javascript" src="jquery.signalR-2.0.3.js"> 1: </script> 2: <script type="text/javascript" src="app.js"></script> 4: </head> 5: <body> 6: <h1>SignalR Auto Disconnect with AngularJS by Shaun</h1> 7: <div> 8: <a href="javascript:void(0)" data-ui-sref="view1">View 1</a> | 9: <a href="javascript:void(0)" data-ui-sref="view2">View 2</a> 10: </div> 11: <div data-ui-view></div> 12: </body> 13: </html> Below is the "app.js". My SignalR logic was in the "View1" page and it will connect to server once the controller was executed. User can specify a user name and send to server, all clients that located in this page will receive the server side greeting message through SignalR. 1: 'use strict'; 2:  3: var app = angular.module('demo', ['ui.router']); 4:  5: app.config(['$stateProvider', '$locationProvider', function ($stateProvider, $locationProvider) { 6: $stateProvider.state('view1', { 7: url: '/view1', 8: templateUrl: 'view1.html', 9: controller: 'View1Ctrl' }); 10:  11: $stateProvider.state('view2', { 12: url: '/view2', 13: templateUrl: 'view2.html', 14: controller: 'View2Ctrl' }); 15:  16: $locationProvider.html5Mode(true); 17: }]); 18:  19: app.value('$', $); 20: app.value('endpoint', 'http://localhost:60448'); 21: app.value('hub', 'GreetingHub'); 22:  23: app.controller('View1Ctrl', function ($scope, $, endpoint, hub) { 24: $scope.user = ''; 25: $scope.response = ''; 26:  27: $scope.greeting = function () { 28: proxy.invoke('Hello', $scope.user) 29: .done(function () {}) 30: .fail(function (error) { 31: console.log(error); 32: }); 33: }; 34:  35: var connection = $.hubConnection(endpoint); 36: var proxy = connection.createHubProxy(hub); 37: proxy.on('hello', function (response) { 38: $scope.$apply(function () { 39: $scope.response = response; 40: }); 41: }); 42: connection.start() 43: .done(function () { 44: console.log('signlar connection established'); 45: }) 46: .fail(function (error) { 47: console.log(error); 48: }); 49: }); 50:  51: app.controller('View2Ctrl', function ($scope, $) { 52: }); When we went to View1 the server side "OnConnect" method will be invoked as below. And in any page we send the message to server, all clients will got the response. If we close one of the client, the server side "OnDisconnect" method will be invoked which is correct. But is we click "View 2" link in the page "OnDisconnect" method will not be invoked even though the content and browser address had been changed. This might cause many SignalR connections remain between the client and server. Below is what happened after I clicked "View 1" and "View 2" links four times. As you can see there are 4 live connections.   Solution Since the reason of this issue is because, AngularJS hijacks the page event that SignalR need to stop the connection, we can handle AngularJS route or state change event and stop SignalR connect manually. In the code below I moved the "connection" variant to global scope, added a handler to "$stateChangeStart" and invoked "stop" method of "connection" if its state was not "disconnected". 1: var connection; 2: app.run(['$rootScope', function ($rootScope) { 3: $rootScope.$on('$stateChangeStart', function () { 4: if (connection && connection.state && connection.state !== 4 /* disconnected */) { 5: console.log('signlar connection abort'); 6: connection.stop(); 7: } 8: }); 9: }]); Now if we refresh the page and navigated to View 1, the connection will be opened. At this state if we clicked "View 2" link the content will be changed and the SignalR connection will be closed automatically.   Summary In this post I demonstrated an issue when we are using SignalR with AngularJS. The connection cannot be closed automatically when we navigate to other page/state in AngularJS. And the solution I mentioned below is to move the SignalR connection as a global variant and close it manually when AngularJS route/state changed. You can download the full sample code here. Moving the SignalR connection as a global variant might not be a best solution. It's just for easy to demo here. In production code I suggest wrapping all SignalR operations into an AngularJS factory. Since AngularJS factory is a singleton object, we can safely put the connection variant in the factory function scope.   Hope this helps, Shaun All documents and related graphics, codes are provided "AS IS" without warranty of any kind. Copyright © Shaun Ziyan Xu. This work is licensed under the Creative Commons License.

    Read the article

  • Connecting Linux to WatchGuard Firebox SSL (OpenVPN client)

    Recently, I got a new project assignment that requires to connect permanently to the customer's network through VPN. They are using a so-called SSL VPN. As I am using OpenVPN since more than 5 years within my company's network I was quite curious about their solution and how it would actually be different from OpenVPN. Well, short version: It is a disguised version of OpenVPN. Unfortunately, the company only offers a client for Windows and Mac OS which shouldn't bother any Linux user after all. OpenVPN is part of every recent distribution and can be activated in a couple of minutes - both client as well as server (if necessary). WatchGuard Firebox SSL - About dialog Borrowing some files from a Windows client installation Initially, I didn't know about the product, so therefore I went through the installation on Windows 8. No obstacles (and no restart despite installation of TAP device drivers!) here and the secured VPN channel was up and running in less than 2 minutes or so. Much appreciated from both parties - customer and me. Of course, this whole client package and my long year approved and stable installation ignited my interest to have a closer look at the WatchGuard client. Compared to the original OpenVPN client (okay, I have to admit this is years ago) this commercial product is smarter in terms of file locations during installation. You'll be able to access the configuration and key files below your roaming application data folder. To get there, simply enter '%AppData%\WatchGuard\Mobile VPN' in your Windows/File Explorer and confirm with Enter/Return. This will display the following files: Application folder below user profile with configuration and certificate files From there we are going to borrow four files, namely: ca.crt client.crt client.ovpn client.pem and transfer them to the Linux system. You might also be able to isolate those four files from a Mac OS client. Frankly, I'm just too lazy to run the WatchGuard client installation on a Mac mini only to find the folder location, and I'm going to describe why a little bit further down this article. I know that you can do that! Feedback in the comment section is appreciated. Configuration of OpenVPN (console) Depending on your distribution the following steps might be a little different but in general you should be able to get the important information from it. I'm going to describe the steps in Ubuntu 13.04 (Raring Ringtail). As usual, there are two possibilities to achieve your goal: console and UI. Let's what it is necessary to be done. First of all, you should ensure that you have OpenVPN installed on your system. Open your favourite terminal application and run the following statement: $ sudo apt-get install openvpn network-manager-openvpn network-manager-openvpn-gnome Just to be on the safe side. The four above mentioned files from your Windows machine could be copied anywhere but either you place them below your own user directory or you put them (as root) below the default directory: /etc/openvpn At this stage you would be able to do a test run already. Just in case, run the following command and check the output (it's the similar information you would get from the 'View Logs...' context menu entry in Windows: $ sudo openvpn --config client.ovpn Pay attention to the correct path to your configuration and certificate files. OpenVPN will ask you to enter your Auth Username and Auth Password in order to establish the VPN connection, same as the Windows client. Remote server and user authentication to establish the VPN Please complete the test run and see whether all went well. You can disconnect pressing Ctrl+C. Simplifying your life - authentication file In my case, I actually set up the OpenVPN client on my gateway/router. This establishes a VPN channel between my network and my client's network and allows me to switch machines easily without having the necessity to install the WatchGuard client on each and every machine. That's also very handy for my various virtualised Windows machines. Anyway, as the client configuration, key and certificate files are located on a headless system somewhere under the roof, it is mandatory to have an automatic connection to the remote site. For that you should first change the file extension '.ovpn' to '.conf' which is the default extension on Linux systems for OpenVPN, and then open the client configuration file in order to extend an existing line. $ sudo mv client.ovpn client.conf $ sudo nano client.conf You should have a similar content to this one here: dev tunclientproto tcp-clientca ca.crtcert client.crtkey client.pemtls-remote "/O=WatchGuard_Technologies/OU=Fireware/CN=Fireware_SSLVPN_Server"remote-cert-eku "TLS Web Server Authentication"remote 1.2.3.4 443persist-keypersist-tunverb 3mute 20keepalive 10 60cipher AES-256-CBCauth SHA1float 1reneg-sec 3660nobindmute-replay-warningsauth-user-pass auth.txt Note: I changed the IP address of the remote directive above (which should be obvious, right?). Anyway, the required change is marked in red and we have to create a new authentication file 'auth.txt'. You can give the directive 'auth-user-pass' any file name you'd like to. Due to my existing OpenVPN infrastructure my setup differs completely from the above written content but for sake of simplicity I just keep it 'as-is'. Okay, let's create this file 'auth.txt' $ sudo nano auth.txt and just put two lines of information in it - username on the first, and password on the second line, like so: myvpnusernameverysecretpassword Store the file, change permissions, and call openvpn with your configuration file again: $ sudo chmod 0600 auth.txt $ sudo openvpn --config client.conf This should now work without being prompted to enter username and password. In case that you placed your files below the system-wide location /etc/openvpn you can operate your VPNs also via service command like so: $ sudo service openvpn start client $ sudo service openvpn stop client Using Network Manager For newer Linux users or the ones with 'console-phobia' I'm going to describe now how to use Network Manager to setup the OpenVPN client. For this move your mouse to the systray area and click on Network Connections => VPN Connections => Configure VPNs... which opens your Network Connections dialog. Alternatively, use the HUD and enter 'Network Connections'. Network connections overview in Ubuntu Click on 'Add' button. On the next dialog select 'Import a saved VPN configuration...' from the dropdown list and click on 'Create...' Choose connection type to import VPN configuration Now you navigate to your folder where you put the client files from the Windows system and you open the 'client.ovpn' file. Next, on the tab 'VPN' proceed with the following steps (directives from the configuration file are referred): General Check the IP address of Gateway ('remote' - we used 1.2.3.4 in this setup) Authentication Change Type to 'Password with Certificates (TLS)' ('auth-pass-user') Enter User name to access your client keys (Auth Name: myvpnusername) Enter Password (Auth Password: verysecretpassword) and choose your password handling Browse for your User Certificate ('cert' - should be pre-selected with client.crt) Browse for your CA Certificate ('ca' - should be filled as ca.crt) Specify your Private Key ('key' - here: client.pem) Then click on the 'Advanced...' button and check the following values: Use custom gateway port: 443 (second value of 'remote' directive) Check the selected value of Cipher ('cipher') Check HMAC Authentication ('auth') Enter the Subject Match: /O=WatchGuard_Technologies/OU=Fireware/CN=Fireware_SSLVPN_Server ('tls-remote') Finally, you have to confirm and close all dialogs. You should be able to establish your OpenVPN-WatchGuard connection via Network Manager. For that, click on the 'VPN Connections => client' entry on your Network Manager in the systray. It is advised that you keep an eye on the syslog to see whether there are any problematic issues that would require some additional attention. Advanced topic: routing As stated above, I'm running the 'WatchGuard client for Linux' on my head-less server, and since then I'm actually establishing a secure communication channel between two networks. In order to enable your network clients to get access to machines on the remote side there are two possibilities to enable that: Proper routing on both sides of the connection which enables both-direction access, or Network masquerading on the 'client side' of the connection Following, I'm going to describe the second option a little bit more in detail. The Linux system that I'm using is already configured as a gateway to the internet. I won't explain the necessary steps to do that, and will only focus on the additional tweaks I had to do. You can find tons of very good instructions and tutorials on 'How to setup a Linux gateway/router' - just use Google. OK, back to the actual modifications. First, we need to have some information about the network topology and IP address range used on the 'other' side. We can get this very easily from /var/log/syslog after we established the OpenVPN channel, like so: $ sudo tail -n20 /var/log/syslog Or if your system is quite busy with logging, like so: $ sudo less /var/log/syslog | grep ovpn The output should contain PUSH received message similar to the following one: Jul 23 23:13:28 ios1 ovpn-client[789]: PUSH: Received control message: 'PUSH_REPLY,topology subnet,route 192.168.1.0 255.255.255.0,dhcp-option DOMAIN ,route-gateway 192.168.6.1,topology subnet,ping 10,ping-restart 60,ifconfig 192.168.6.2 255.255.255.0' The interesting part for us is the route command which I highlighted already in the sample PUSH_REPLY. Depending on your remote server there might be multiple networks defined (172.16.x.x and/or 10.x.x.x). Important: The IP address range on both sides of the connection has to be different, otherwise you will have to shuffle IPs or increase your the netmask. {loadposition content_adsense} After the VPN connection is established, we have to extend the rules for iptables in order to route and masquerade IP packets properly. I created a shell script to take care of those steps: #!/bin/sh -eIPTABLES=/sbin/iptablesDEV_LAN=eth0DEV_VPNS=tun+VPN=192.168.1.0/24 $IPTABLES -A FORWARD -i $DEV_LAN -o $DEV_VPNS -d $VPN -j ACCEPT$IPTABLES -A FORWARD -i $DEV_VPNS -o $DEV_LAN -s $VPN -j ACCEPT$IPTABLES -t nat -A POSTROUTING -o $DEV_VPNS -d $VPN -j MASQUERADE I'm using the wildcard interface 'tun+' because I have multiple client configurations for OpenVPN on my server. In your case, it might be sufficient to specify device 'tun0' only. Simplifying your life - automatic connect on boot Now, that the client connection works flawless, configuration of routing and iptables is okay, we might consider to add another 'laziness' factor into our setup. Due to kernel updates or other circumstances it might be necessary to reboot your system. Wouldn't it be nice that the VPN connections are established during the boot procedure? Yes, of course it would be. To achieve this, we have to configure OpenVPN to automatically start our VPNs via init script. Let's have a look at the responsible 'default' file and adjust the settings accordingly. $ sudo nano /etc/default/openvpn Which should have a similar content to this: # This is the configuration file for /etc/init.d/openvpn## Start only these VPNs automatically via init script.# Allowed values are "all", "none" or space separated list of# names of the VPNs. If empty, "all" is assumed.# The VPN name refers to the VPN configutation file name.# i.e. "home" would be /etc/openvpn/home.conf#AUTOSTART="all"#AUTOSTART="none"#AUTOSTART="home office"## ... more information which remains unmodified ... With the OpenVPN client configuration as described above you would either set AUTOSTART to "all" or to "client" to enable automatic start of your VPN(s) during boot. You should also take care that your iptables commands are executed after the link has been established, too. You can easily test this configuration without reboot, like so: $ sudo service openvpn restart Enjoy stable VPN connections between your Linux system(s) and a WatchGuard Firebox SSL remote server. Cheers, JoKi

    Read the article

  • How do I align my partition table properly?

    - by Jorge Castro
    I am in the process of building my first RAID5 array. I've used mdadm to create the following set up: root@bondigas:~# mdadm --detail /dev/md1 /dev/md1: Version : 00.90 Creation Time : Wed Oct 20 20:00:41 2010 Raid Level : raid5 Array Size : 5860543488 (5589.05 GiB 6001.20 GB) Used Dev Size : 1953514496 (1863.02 GiB 2000.40 GB) Raid Devices : 4 Total Devices : 4 Preferred Minor : 1 Persistence : Superblock is persistent Update Time : Wed Oct 20 20:13:48 2010 State : clean, degraded, recovering Active Devices : 3 Working Devices : 4 Failed Devices : 0 Spare Devices : 1 Layout : left-symmetric Chunk Size : 64K Rebuild Status : 1% complete UUID : f6dc829e:aa29b476:edd1ef19:85032322 (local to host bondigas) Events : 0.12 Number Major Minor RaidDevice State 0 8 16 0 active sync /dev/sdb 1 8 32 1 active sync /dev/sdc 2 8 48 2 active sync /dev/sdd 4 8 64 3 spare rebuilding /dev/sde While that's going I decided to format the beast with the following command: root@bondigas:~# mkfs.ext4 /dev/md1p1 mke2fs 1.41.11 (14-Mar-2010) /dev/md1p1 alignment is offset by 63488 bytes. This may result in very poor performance, (re)-partitioning suggested. Filesystem label= OS type: Linux Block size=4096 (log=2) Fragment size=4096 (log=2) Stride=16 blocks, Stripe width=48 blocks 97853440 inodes, 391394047 blocks 19569702 blocks (5.00%) reserved for the super user First data block=0 Maximum filesystem blocks=0 11945 block groups 32768 blocks per group, 32768 fragments per group 8192 inodes per group Superblock backups stored on blocks: 32768, 98304, 163840, 229376, 294912, 819200, 884736, 1605632, 2654208, 4096000, 7962624, 11239424, 20480000, 23887872, 71663616, 78675968, 102400000, 214990848 Writing inode tables: ^C 27/11945 root@bondigas:~# ^C I am unsure what to do about "/dev/md1p1 alignment is offset by 63488 bytes." and how to properly partition the disks to match so I can format it properly.

    Read the article

  • Oracle Desktop Virtualization at HIMSS 2011

    - by chris.kawalek(at)oracle.com
    The HIMSS Conference is an extremely important industry trade show put on by The Healthcare Information and Management Systems Society. It's being held in Florida starting this Sunday, February 20th. Their slogan, "Linking people, potential, and progress" could be true of Oracle desktop virtualization as well! The Oracle desktop virtualization group has worked very closely with the Oracle healthcare business unit to have a large presence at this show, and I wanted to tell you a bit about what we're doing: - All Oracle demos are being done on Sun Ray Clients That's right, every demo pod in the large Oracle booth will have a Sun Ray Client with each demo tied to a smart card. Too many people at your demo station? Pop your card out and go to a different one. We'll also be demoing Oracle desktop virtualization at a dedicated demo station, too. This is great stuff! Find Oracle at booth #1651 Oracle's page about HIMSS - Focus Group - Caregiver Mobility with Oracle Sun Ray Clients and Desktop Virtualization Feb 22, 3:15-4:15 PM This focus group will be for customers interested in Oracle desktop virtualization. It's invitation only, but you can comment on this blog post and we can give you info on how to attend (your comment won't be made public). - Solution Session - Fast, Secure, Workflow Optimized: Inexpensive Access to Care Information is Possible Inside and Outside of the Hospital Feb 23, 4:15 PM Booth #685, Wireless and Mobility Theatre Oracle's Adam Workman will cover caregiver mobility and the benefits of Oracle desktop virtualization to healthcare organizations. - New healthcare solutions page on oracle.com We've created a page dedicated to content involving desktop virtualization and healthcare. This will be your onestop shop if looking for desktop virtualization and healthcare information. - New desktop virtualization and healthcare solution data sheet This document outlines how we define "Caregiver Mobility" and how Oracle products are used to facilitate quicker, more secure access to patient data. We'll have some more updates from the show next week. It looks like its going to be an exciting event! -Chris

    Read the article

  • Set umask, set permissions, and set ACL, but SAMBA isn't using those?

    - by Kris Anderson
    I'm running on Ubuntu Server 12.04. I have a folder called Music and I want the default folder permissions to be 775 and the default file to then be 664. I set the default permissions on the Music folder to be 775. I configured ACL to use these default permissions as well: file: Music owner: kris group: kris flags: ss- user::rwx group::rwx other::r-x default:user::rwx default:group::rwx default:other::r-x I also changed the default umask for my user account, kris, to 002 in .profile. Shouldn't and new file/folder now use those permissions when writing to the Samba share? ACL should work with Samba from what I can gather. Currently, if I write to that folder using my mac, folders are getting 755 and files 644. I have another app on my mac called GoodSync which which is able to sync a local directory on my mac to a network samba share, but those permissions are even worse. files are being written as 700 using that program. So it looks like Samba is allowing the host/program to determine the folder/file permissions. What changes do I need to make to force the permissions I want regardless of what the host tries to write on the server?

    Read the article

  • 3 Key Trends For Mobile Commerce – Location, Location, Location

    - by Michael Hylton
    This past weekend I was at a major bookstore chain and looking for a particular book.  Rather than ask the clerk, I went to my smartphone and went online to find the book title, author, and competing price.  I know I’m not alone in this effort and more and more individuals (and businesses) will use the power of mobility to tilt the scale in their favor. Armed with a mobile device – smartphone or tablet – folks will use them to research, compare, and ultimately purchase.  A recent PayPal survey found that 46% of respondents plan to use a mobile device this holiday season to make a purchase.   An astounding 27% of consumers in an e-tailing group survey commissioned by Oracle, use a tablet device daily or several times a week to research products and services. Beyond researching or making purchases, 35% of consumers use their smartphone to receive offers and coupons, and 32% access coupons and redeem them at their local retail store.  And with GPS capabilities in smartphones and tablet (and with user’s approval), retailers will start pushing coupons and offers directly to phone users based on their proximity to their store (or their competitors). Security is one concern that both shoppers, companies and phone manufacturers will have to deal with in the coming years.  In that same Oracle-sponsored e-tailing group consumer survey, 32% of consumers were concerned about giving their credit card information via a smartphone. You can gain further insight into the mind of today’s consumer by reading the e-tailing group white paper, titled “the connected consumer”.

    Read the article

  • 3 Key Trends For Mobile Commerce – Location, Location, Location

    - by Michael Hylton
    This past weekend I was at a major bookstore chain and looking for a particular book.  Rather than ask the clerk, I went to my smartphone and went online to find the book title, author, and competing price.  I know I’m not alone in this effort and more and more individuals (and businesses) will use the power of mobility to tilt the scale in their favor. Armed with a mobile device – smartphone or tablet – folks will use them to research, compare, and ultimately purchase.  A recent PayPal survey found that 46% of respondents plan to use a mobile device this holiday season to make a purchase.   An astounding 27% of consumers in an e-tailing group survey commissioned by Oracle, use a tablet device daily or several times a week to research products and services. Beyond researching or making purchases, 35% of consumers use their smartphone to receive offers and coupons, and 32% access coupons and redeem them at their local retail store.  And with GPS capabilities in smartphones and tablet (and with user’s approval), retailers will start pushing coupons and offers directly to phone users based on their proximity to their store (or their competitors). Security is one concern that both shoppers, companies and phone manufacturers will have to deal with in the coming years.  In that same Oracle-sponsored e-tailing group consumer survey, 32% of consumers were concerned about giving their credit card information via a smartphone. You can gain further insight into the mind of today’s consumer by reading the e-tailing group white paper, titled “the connected consumer”.

    Read the article

  • Visual Studio ALM MVP of the Year 2011

    - by Martin Hinshelwood
    For some reason this year some of my peers decided to vote for me as a contender for Visual Studio ALM MVP of the year. I am not sure what I did to deserve this, but a number of people have commented that I have a rather useful blog. I feel wholly unworthy to join the ranks of previous winners: Ed Blankenship (2010) Martin Woodward (2009) Thank you to everyone who voted regardless of who you voted for. If there was a prize for the best group of MVP’s then the Visual Studio ALM MVP would be a clear winner, as would the product group of product groups that is Visual Studio ALM Group. To use a phrase that I have learned since moving to Seattle and probably use too much: you guys are all just awesome. I have tried my best in the last year to document not only every problem that I have had with Team Foundation Server (TFS), but also to document as many of the things I am doing as possible. I have taken some of Adam Cogan’s rules to heart and when a customer asks me a question I always blog the answer and send them a link. This allows both my blog and my understanding of TFS to grow while creating a useful bank of content. The idea is that if one customer asks, all benefit. I try, when writing for my blog, to capture both the essence and the context for a problem being solved. This allows more people to benefit as they do not need to understand the specifics of an environment to gain value. I have a number of goals for this year that I think will help increase value in the community: persuade my new colleagues at Northwest Cadence to do more blogging (Steve, Jeff, Shad and Rennie) Rangers Project – TFS Iteration Automation with Willy-Peter Schaub, Bill Essary, Martin Hinshelwood, Mike Fourie, Jeff Bramwell and Brian Blackman Write a book on the Team Foundation Server API with Willy-Peter Schaub, Mike Fourie and Jeff Bramwell write more useful blog posts I do not think that these things are beyond the realms of do-ability, but we will see…

    Read the article

  • Click Once Deployment Process and Issue Resolution

    - by Geordie
    Introduction We are adopting Click Once as a deployment standard for Thick .Net application clients.  The latest version of this tool has matured it to a point where it can be used in an enterprise environment.  This guide will identify how to use Click Once deployment and promote code trough the dev, test and production environments. Why Use Click Once over SCCM If we already use SCCM why add Click Once to the deployment options.  The advantages of Click Once are their ability to update the code in a single location and have the update flow automatically down to the user community.  There have been challenges in the past with getting configuration updates to download but these can now be achieved.  With SCCM you can do the same thing but it then needs to be packages and pushed out to users.  Each time a new user is added to an application, time needs to be spent by an administrator, to push out any required application packages.  With Click Once the user would go to a web link and the application and pre requisites will automatically get installed. New Deployment Steps Overview The deployment in an enterprise environment includes several steps as the solution moves through the development life cycle before being released into production.  To make mitigate risk during the release phase, it is important to ensure the solution is not deployed directly into production from the development tools.  Although this is the easiest path, it can introduce untested code into production and result in unexpected results. 1. Deploy the client application to a development web server using Visual Studio 2008 Click Once deployment tools.  Once potential production versions of the solution are being generated, ensure the production install URL is specified when deploying code from Visual Studio.  (For details see ‘Deploying Click Once Code from Visual Studio’) 2. xCopy the code to the test server.  Run the MageUI tool to update the URLs, signing and version numbers to match the test server. (For details see ‘Moving Click Once Code to a new Server without using Visual Studio’) 3. xCopy the code to the production server.  Run the MageUI tool to update the URLs, signing and version numbers to match the production server. The certificate used to sign the code should be provided by a certificate authority that will be trusted by the client machines.  Finally make sure the setup.exe contains the production install URL.  If not redeploy the solution from Visual Studio to the dev environment specifying the production install URL.  Then xcopy the install.exe file from dev to production.  (For details see ‘Moving Click Once Code to a new Server without using Visual Studio’) Detailed Deployment Steps Deploying Click Once Code From Visual Studio Open Visual Studio and create a new WinForms or WPF project.   In the solution explorer right click on the project and select ‘Publish’ in the context menu.   The ‘Publish Wizard’ will start.  Enter the development deployment path.  This could be a local directory or web site.  When first publishing the solution set this to a development web site and Visual basic will create a site with an install.htm page.  Click Next.  Select weather the application will be available both online and offline. Then click Finish. Once the initial deployment is completed, republish the solution this time mapping to the directory that holds the code that was just published.  This time the Publish Wizard contains and additional option.   The setup.exe file that is created has the install URL hardcoded in it.  It is this screen that allows you to specify the URL to use.  At some point a setup.exe file must be generated for production.  Enter the production URL and deploy the solution to the dev folder.  This file can then be saved for latter use in deployment to production.  During development this URL should be pointing to development site to avoid accidently installing the production application. Visual studio will publish the application to the desired location in the process it will create an anonymous ‘pfx’ certificate to sign the deployment configuration files.  A production certificate should be acquired in preparation for deployment to production.   Directory structure created by Visual Studio     Application files created by Visual Studio   Development web site (install.htm) created by Visual Studio Migrating Click Once Code to a new Server without using Visual Studio To migrate the Click Once application code to a new server, a tool called MageUI is needed to modify the .application and .manifest files.  The MageUI tool is usually located – ‘C:\Program Files\Microsoft SDKs\Windows\v6.0A\Bin’ folder or can be downloaded from the web. When deploying to a new environment copy all files in the project folder to the new server.  In this case the ‘ClickOnceSample’ folder and contents.  The old application versions can be deleted, in this case ‘ClickOnceSample_1_0_0_0’ and ‘ClickOnceSample_1_0_0_1’.  Open IIS Manager and create a virtual directory that points to the project folder.  Also make the publish.htm the default web page.   Run the ManeUI tool and then open the .application file in the root project folder (in this case in the ‘ClickOnceSample’ folder). Click on the Deployment Options in the left hand list and update the URL to the new server URL and save the changes.   When MageUI tries to save the file it will prompt for the file to be signed.   This step cannot be bypassed if you want the Click Once deployment to work from a web site.  The easiest solution to this for test is to use the auto generated certificate that Visual Studio created for the project.  This certificate can be found with the project source code.   To save time go to File>Preferences and configure the ‘Use default signing certificate’ fields.   Future deployments will only require application files to be transferred to the new server.  The only difference is then updating the .application file the ‘Version’ must be updated to match the new version and the ‘Application Reference’ has to be update to point to the new .manifest file.     Updating the Configuration File of a Click Once Deployment Package without using Visual Studio When an update to the configuration file is required, modifying the ClickOnceSample.exe.config.deploy file will not result in current users getting the new configurations.  We do not want to go back to Visual Studio and generate a new version as this might introduce unexpected code changes.  A new version of the application can be created by copying the folder (in this case ClickOnceSample_1_0_0_2) and pasting it into the application Files directory.  Rename the directory ‘ClickOnceSample_1_0_0_3’.  In the new folder open the configuration file in notepad and make the configuration changes. Run MageUI and open the manifest file in the newly copied directory (ClickOnceSample_1_0_0_3).   Edit the manifest version to reflect the newly copied files (in this case 1.0.0.3).  Then save the file.  Open the .application file in the root folder.  Again update the version to 1.0.0.3.  Since the file has not changed the Deployment Options/Start Location URL should still be correct.  The application Reference needs to be updated to point to the new versions .manifest file.  Save the file. Next time a user runs the application the new version of the configuration file will be down loaded.  It is worth noting that there are 2 different types of configuration parameter; application and user.  With Click Once deployment the difference is significant.  When an application is downloaded the configuration file is also brought down to the client machine.  The developer may have written code to update the user parameters in the application.  As a result each time a new version of the application is down loaded the user parameters are at risk of being overwritten.  With Click Once deployment the system knows if the user parameters are still the default values.  If they are they will be overwritten with the new default values in the configuration file.  If they have been updated by the user, they will not be overwritten. Settings configuration view in Visual Studio Production Deployment When deploying the code to production it is prudent to disable the development and test deployment sites.  This will allow errors such as incorrect URL to be quickly identified in the initial testing after deployment.  If the sites are active there is no way to know if the application was downloaded from the production deployment and not redirected to test or dev.   Troubleshooting Clicking the install button on the install.htm page fails. Error: URLDownloadToCacheFile failed with HRESULT '-2146697210' Error: An error occurred trying to download <file>   This is due to the setup.exe file pointing to the wrong location. ‘The setup.exe file that is created has the install URL hardcoded in it.  It is this screen that allows you to specify the URL to use.  At some point a setup.exe file must be generated for production.  Enter the production URL and deploy the solution to the dev folder.  This file can then be saved for latter use in deployment to production.  During development this URL should be pointing to development site to avoid accidently installing the production application.’

    Read the article

  • Salary and profit distribution in game industry?

    - by drowneath
    A couple years ago, I started a group/team of passionate people in game development. I was the one who had the idea to form a group that will (hopefully) later be a company/real studio. I was the one who gathered the people too. We are consisting of only a few people (< 10 people) and everyone has their own specialties in game development. For some reason, everyone agreed to make me the executive director of the group. We are currently focused in creating flash games and mobile games. Until now, we have created a few free game titles and gained profit from some freelancing projects. Since I have no prior experience in running a "company", I decided to split the profit we gained from projects equally regardless of the member's role in the company, as long as he/she is involved in and have contributed a decent amount of work to the development of the project. My questions are: What is the correct way to split profit that is gained from freelance projects that are developed together? Once we've released enough products and ready to register our company legally, what about the salary? What benefits do I have from being the founder and the director? I'm not a control-freak, but I want everything to be clear.

    Read the article

< Previous Page | 520 521 522 523 524 525 526 527 528 529 530 531  | Next Page >