Search Results

Search found 41230 results on 1650 pages for 'type safety'.

Page 976/1650 | < Previous Page | 972 973 974 975 976 977 978 979 980 981 982 983  | Next Page >

• Problem with setup VPN on Ubuntu Server 12.04

  - by Yozone W.

  I have a problem with setup VPN server on my Ubuntu VPS, here is my server environments: Ubuntu Server 12.04 x86_64 xl2tpd 1.3.1+dfsg-1 pppd 2.4.5-5ubuntu1 openswan 1:2.6.38-1~precise1 After install software and configuration: ipsec verify Checking your system to see if IPsec got installed and started correctly: Version check and ipsec on-path [OK] Linux Openswan U2.6.38/K3.2.0-24-virtual (netkey) Checking for IPsec support in kernel [OK] SAref kernel support [N/A] NETKEY: Testing XFRM related proc values [OK] [OK] [OK] Checking that pluto is running [OK] Pluto listening for IKE on udp 500 [OK] Pluto listening for NAT-T on udp 4500 [OK] Checking for 'ip' command [OK] Checking /bin/sh is not /bin/dash [WARNING] Checking for 'iptables' command [OK] Opportunistic Encryption Support [DISABLED] /var/log/auth.log message: Oct 16 06:50:54 vpn pluto[3963]: packet from [My IP Address]:2251: received Vendor ID payload [RFC 3947] method set to=115 Oct 16 06:50:54 vpn pluto[3963]: packet from [My IP Address]:2251: received Vendor ID payload [draft-ietf-ipsec-nat-t-ike] meth=114, but already using method 115 Oct 16 06:50:54 vpn pluto[3963]: packet from [My IP Address]:2251: received Vendor ID payload [draft-ietf-ipsec-nat-t-ike-08] meth=113, but already using method 115 Oct 16 06:50:54 vpn pluto[3963]: packet from [My IP Address]:2251: received Vendor ID payload [draft-ietf-ipsec-nat-t-ike-07] meth=112, but already using method 115 Oct 16 06:50:54 vpn pluto[3963]: packet from [My IP Address]:2251: received Vendor ID payload [draft-ietf-ipsec-nat-t-ike-06] meth=111, but already using method 115 Oct 16 06:50:54 vpn pluto[3963]: packet from [My IP Address]:2251: received Vendor ID payload [draft-ietf-ipsec-nat-t-ike-05] meth=110, but already using method 115 Oct 16 06:50:54 vpn pluto[3963]: packet from [My IP Address]:2251: received Vendor ID payload [draft-ietf-ipsec-nat-t-ike-04] meth=109, but already using method 115 Oct 16 06:50:54 vpn pluto[3963]: packet from [My IP Address]:2251: received Vendor ID payload [draft-ietf-ipsec-nat-t-ike-03] meth=108, but already using method 115 Oct 16 06:50:54 vpn pluto[3963]: packet from [My IP Address]:2251: received Vendor ID payload [draft-ietf-ipsec-nat-t-ike-02] meth=107, but already using method 115 Oct 16 06:50:54 vpn pluto[3963]: packet from [My IP Address]:2251: received Vendor ID payload [draft-ietf-ipsec-nat-t-ike-02_n] meth=106, but already using method 115 Oct 16 06:50:54 vpn pluto[3963]: packet from [My IP Address]:2251: ignoring Vendor ID payload [FRAGMENTATION 80000000] Oct 16 06:50:54 vpn pluto[3963]: packet from [My IP Address]:2251: received Vendor ID payload [Dead Peer Detection] Oct 16 06:50:54 vpn pluto[3963]: "L2TP-PSK-NAT"[5] [My IP Address] #5: responding to Main Mode from unknown peer [My IP Address] Oct 16 06:50:54 vpn pluto[3963]: "L2TP-PSK-NAT"[5] [My IP Address] #5: transition from state STATE_MAIN_R0 to state STATE_MAIN_R1 Oct 16 06:50:54 vpn pluto[3963]: "L2TP-PSK-NAT"[5] [My IP Address] #5: STATE_MAIN_R1: sent MR1, expecting MI2 Oct 16 06:50:55 vpn pluto[3963]: "L2TP-PSK-NAT"[5] [My IP Address] #5: NAT-Traversal: Result using draft-ietf-ipsec-nat-t-ike (MacOS X): peer is NATed Oct 16 06:50:55 vpn pluto[3963]: "L2TP-PSK-NAT"[5] [My IP Address] #5: transition from state STATE_MAIN_R1 to state STATE_MAIN_R2 Oct 16 06:50:55 vpn pluto[3963]: "L2TP-PSK-NAT"[5] [My IP Address] #5: STATE_MAIN_R2: sent MR2, expecting MI3 Oct 16 06:50:55 vpn pluto[3963]: "L2TP-PSK-NAT"[5] [My IP Address] #5: ignoring informational payload, type IPSEC_INITIAL_CONTACT msgid=00000000 Oct 16 06:50:55 vpn pluto[3963]: "L2TP-PSK-NAT"[5] [My IP Address] #5: Main mode peer ID is ID_IPV4_ADDR: '192.168.12.52' Oct 16 06:50:55 vpn pluto[3963]: "L2TP-PSK-NAT"[5] [My IP Address] #5: switched from "L2TP-PSK-NAT" to "L2TP-PSK-NAT" Oct 16 06:50:55 vpn pluto[3963]: "L2TP-PSK-NAT"[6] [My IP Address] #5: deleting connection "L2TP-PSK-NAT" instance with peer [My IP Address] {isakmp=#0/ipsec=#0} Oct 16 06:50:55 vpn pluto[3963]: "L2TP-PSK-NAT"[6] [My IP Address] #5: transition from state STATE_MAIN_R2 to state STATE_MAIN_R3 Oct 16 06:50:55 vpn pluto[3963]: "L2TP-PSK-NAT"[6] [My IP Address] #5: new NAT mapping for #5, was [My IP Address]:2251, now [My IP Address]:2847 Oct 16 06:50:55 vpn pluto[3963]: "L2TP-PSK-NAT"[6] [My IP Address] #5: STATE_MAIN_R3: sent MR3, ISAKMP SA established {auth=OAKLEY_PRESHARED_KEY cipher=aes_256 prf=oakley_sha group=modp1024} Oct 16 06:50:55 vpn pluto[3963]: "L2TP-PSK-NAT"[6] [My IP Address] #5: Dead Peer Detection (RFC 3706): enabled Oct 16 06:50:56 vpn pluto[3963]: "L2TP-PSK-NAT"[6] [My IP Address] #5: the peer proposed: [My Server IP Address]/32:17/1701 -> 192.168.12.52/32:17/0 Oct 16 06:50:56 vpn pluto[3963]: "L2TP-PSK-NAT"[6] [My IP Address] #5: NAT-Traversal: received 2 NAT-OA. using first, ignoring others Oct 16 06:50:56 vpn pluto[3963]: "L2TP-PSK-NAT"[6] [My IP Address] #6: responding to Quick Mode proposal {msgid:8579b1fb} Oct 16 06:50:56 vpn pluto[3963]: "L2TP-PSK-NAT"[6] [My IP Address] #6: us: [My Server IP Address]<[My Server IP Address]>:17/1701 Oct 16 06:50:56 vpn pluto[3963]: "L2TP-PSK-NAT"[6] [My IP Address] #6: them: [My IP Address][192.168.12.52]:17/65280===192.168.12.52/32 Oct 16 06:50:56 vpn pluto[3963]: "L2TP-PSK-NAT"[6] [My IP Address] #6: transition from state STATE_QUICK_R0 to state STATE_QUICK_R1 Oct 16 06:50:56 vpn pluto[3963]: "L2TP-PSK-NAT"[6] [My IP Address] #6: STATE_QUICK_R1: sent QR1, inbound IPsec SA installed, expecting QI2 Oct 16 06:50:56 vpn pluto[3963]: "L2TP-PSK-NAT"[6] [My IP Address] #6: Dead Peer Detection (RFC 3706): enabled Oct 16 06:50:56 vpn pluto[3963]: "L2TP-PSK-NAT"[6] [My IP Address] #6: transition from state STATE_QUICK_R1 to state STATE_QUICK_R2 Oct 16 06:50:56 vpn pluto[3963]: "L2TP-PSK-NAT"[6] [My IP Address] #6: STATE_QUICK_R2: IPsec SA established transport mode {ESP=>0x08bda158 <0x4920a374 xfrm=AES_256-HMAC_SHA1 NATOA=192.168.12.52 NATD=[My IP Address]:2847 DPD=enabled} Oct 16 06:51:16 vpn pluto[3963]: "L2TP-PSK-NAT"[6] [My IP Address] #5: received Delete SA(0x08bda158) payload: deleting IPSEC State #6 Oct 16 06:51:16 vpn pluto[3963]: "L2TP-PSK-NAT"[6] [My IP Address] #5: ERROR: netlink XFRM_MSG_DELPOLICY response for flow eroute_connection delete included errno 2: No such file or directory Oct 16 06:51:16 vpn pluto[3963]: "L2TP-PSK-NAT"[6] [My IP Address] #5: received and ignored informational message Oct 16 06:51:16 vpn pluto[3963]: "L2TP-PSK-NAT"[6] [My IP Address] #5: received Delete SA payload: deleting ISAKMP State #5 Oct 16 06:51:16 vpn pluto[3963]: "L2TP-PSK-NAT"[6] [My IP Address]: deleting connection "L2TP-PSK-NAT" instance with peer [My IP Address] {isakmp=#0/ipsec=#0} Oct 16 06:51:16 vpn pluto[3963]: packet from [My IP Address]:2847: received and ignored informational message xl2tpd -D message: xl2tpd[4289]: Enabling IPsec SAref processing for L2TP transport mode SAs xl2tpd[4289]: IPsec SAref does not work with L2TP kernel mode yet, enabling forceuserspace=yes xl2tpd[4289]: setsockopt recvref[30]: Protocol not available xl2tpd[4289]: This binary does not support kernel L2TP. xl2tpd[4289]: xl2tpd version xl2tpd-1.3.1 started on vpn.netools.me PID:4289 xl2tpd[4289]: Written by Mark Spencer, Copyright (C) 1998, Adtran, Inc. xl2tpd[4289]: Forked by Scott Balmos and David Stipp, (C) 2001 xl2tpd[4289]: Inherited by Jeff McAdams, (C) 2002 xl2tpd[4289]: Forked again by Xelerance (www.xelerance.com) (C) 2006 xl2tpd[4289]: Listening on IP address [My Server IP Address], port 1701 Then it just stopped here, and have no any response. I can't connect VPN on my mac client, the /var/log/system.log message: Oct 16 15:17:36 azone-iMac.local configd[17]: SCNC: start, triggered by SystemUIServer, type L2TP, status 0 Oct 16 15:17:36 azone-iMac.local pppd[3799]: pppd 2.4.2 (Apple version 596.13) started by azone, uid 501 Oct 16 15:17:38 azone-iMac.local pppd[3799]: L2TP connecting to server 'vpn.netools.me' ([My Server IP Address])... Oct 16 15:17:38 azone-iMac.local pppd[3799]: IPSec connection started Oct 16 15:17:38 azone-iMac.local racoon[359]: Connecting. Oct 16 15:17:38 azone-iMac.local racoon[359]: IPSec Phase1 started (Initiated by me). Oct 16 15:17:38 azone-iMac.local racoon[359]: IKE Packet: transmit success. (Initiator, Main-Mode message 1). Oct 16 15:17:38 azone-iMac.local racoon[359]: IKE Packet: receive success. (Initiator, Main-Mode message 2). Oct 16 15:17:38 azone-iMac.local racoon[359]: IKE Packet: transmit success. (Initiator, Main-Mode message 3). Oct 16 15:17:38 azone-iMac.local racoon[359]: IKE Packet: receive success. (Initiator, Main-Mode message 4). Oct 16 15:17:38 azone-iMac.local racoon[359]: IKE Packet: transmit success. (Initiator, Main-Mode message 5). Oct 16 15:17:38 azone-iMac.local racoon[359]: IKEv1 Phase1 AUTH: success. (Initiator, Main-Mode Message 6). Oct 16 15:17:38 azone-iMac.local racoon[359]: IKE Packet: receive success. (Initiator, Main-Mode message 6). Oct 16 15:17:38 azone-iMac.local racoon[359]: IKEv1 Phase1 Initiator: success. (Initiator, Main-Mode). Oct 16 15:17:38 azone-iMac.local racoon[359]: IPSec Phase1 established (Initiated by me). Oct 16 15:17:39 azone-iMac.local racoon[359]: IPSec Phase2 started (Initiated by me). Oct 16 15:17:39 azone-iMac.local racoon[359]: IKE Packet: transmit success. (Initiator, Quick-Mode message 1). Oct 16 15:17:39 azone-iMac.local racoon[359]: IKE Packet: receive success. (Initiator, Quick-Mode message 2). Oct 16 15:17:39 azone-iMac.local racoon[359]: IKE Packet: transmit success. (Initiator, Quick-Mode message 3). Oct 16 15:17:39 azone-iMac.local racoon[359]: IKEv1 Phase2 Initiator: success. (Initiator, Quick-Mode). Oct 16 15:17:39 azone-iMac.local racoon[359]: IPSec Phase2 established (Initiated by me). Oct 16 15:17:39 azone-iMac.local pppd[3799]: IPSec connection established Oct 16 15:17:59 azone-iMac.local pppd[3799]: L2TP cannot connect to the server Oct 16 15:17:59 azone-iMac.local racoon[359]: IPSec disconnecting from server [My Server IP Address] Oct 16 15:17:59 azone-iMac.local racoon[359]: IKE Packet: transmit success. (Information message). Oct 16 15:17:59 azone-iMac.local racoon[359]: IKEv1 Information-Notice: transmit success. (Delete IPSEC-SA). Oct 16 15:17:59 azone-iMac.local racoon[359]: IKE Packet: transmit success. (Information message). Oct 16 15:17:59 azone-iMac.local racoon[359]: IKEv1 Information-Notice: transmit success. (Delete ISAKMP-SA). Anyone help? Thanks a million!

  Read the article

• add_shown & add_hiding ModalPopupExtender Events

  - by Yousef_Jadallah

      In this topic, I’ll discuss the Client events we usually need while using ModalPopupExtender. The add_shown fires when the ModalPopupExtender had shown and add_hiding fires when the user cancels it by CancelControlID,note that it fires before hiding the modal. They are useful in many cases, for example may you need to set focus to specific Textbox when the user display the modal, or if you need to reset the controls values inside the Modal after it has been hidden. To declare Client event either in pageLoad javascript function or you can attach the function by Sys.Application.add_load like this: Sys.Application.add_load(modalInit); function modalInit() { var modalPopup = $find('mpeID'); modalPopup.add_hiding(onHiding); } function onHiding(sender, args) { } .csharpcode, .csharpcode pre { font-size: small; color: black; font-family: consolas, "Courier New", courier, monospace; background-color: #ffffff; /*white-space: pre;*/ } .csharpcode pre { margin: 0em; } .csharpcode .rem { color: #008000; } .csharpcode .kwrd { color: #0000ff; } .csharpcode .str { color: #006080; } .csharpcode .op { color: #0000c0; } .csharpcode .preproc { color: #cc6633; } .csharpcode .asp { background-color: #ffff00; } .csharpcode .html { color: #800000; } .csharpcode .attr { color: #ff0000; } .csharpcode .alt { background-color: #f4f4f4; width: 100%; margin: 0em; } .csharpcode .lnum { color: #606060; }   I’ll use the first way in the current example. So lets start with the illustration:   1- In this example am using simple panel which contain UserName and Password Textboxes besides submit and cancel buttons, this Panel will be used as PopupControlID in the ModalPopupExtender : <asp:Panel ID="panModal" runat="server" Height="180px" Width="300px" style="display:none" CssClass="ModalWindow"> <table width="100%" > <tr> <td> User Name </td> <td> <asp:TextBox ID="txtName" runat="server"></asp:TextBox> </td> </tr> <tr> <td> Password </td> <td> <asp:TextBox ID="txtPassword" runat="server" TextMode="Password"></asp:TextBox> </td> </tr> </table> <br /> <asp:Button ID="btnSubmit" runat="server" Text="Submit" /> <asp:Button ID="btnCancel" runat="server" Text="Cancel" /> </asp:Panel>   You can use this simple style for the Panel : <style type="text/css"> .ModalWindow { border: solid; border-width:3px; background:#f0f0f0; } </style> .csharpcode, .csharpcode pre { font-size: small; color: black; font-family: consolas, "Courier New", courier, monospace; background-color: #ffffff; /*white-space: pre;*/ } .csharpcode pre { margin: 0em; } .csharpcode .rem { color: #008000; } .csharpcode .kwrd { color: #0000ff; } .csharpcode .str { color: #006080; } .csharpcode .op { color: #0000c0; } .csharpcode .preproc { color: #cc6633; } .csharpcode .asp { background-color: #ffff00; } .csharpcode .html { color: #800000; } .csharpcode .attr { color: #ff0000; } .csharpcode .alt { background-color: #f4f4f4; width: 100%; margin: 0em; } .csharpcode .lnum { color: #606060; }   2- Create the view button (TargetControlID) as you know this contain the ID of the element that activates the modal popup: <asp:Button ID="btnView" runat="server" Text="View" /> .csharpcode, .csharpcode pre { font-size: small; color: black; font-family: consolas, "Courier New", courier, monospace; background-color: #ffffff; /*white-space: pre;*/ } .csharpcode pre { margin: 0em; } .csharpcode .rem { color: #008000; } .csharpcode .kwrd { color: #0000ff; } .csharpcode .str { color: #006080; } .csharpcode .op { color: #0000c0; } .csharpcode .preproc { color: #cc6633; } .csharpcode .asp { background-color: #ffff00; } .csharpcode .html { color: #800000; } .csharpcode .attr { color: #ff0000; } .csharpcode .alt { background-color: #f4f4f4; width: 100%; margin: 0em; } .csharpcode .lnum { color: #606060; } .csharpcode, .csharpcode pre { font-size: small; color: black; font-family: consolas, "Courier New", courier, monospace; background-color: #ffffff; /*white-space: pre;*/ } .csharpcode pre { margin: 0em; } .csharpcode .rem { color: #008000; } .csharpcode .kwrd { color: #0000ff; } .csharpcode .str { color: #006080; } .csharpcode .op { color: #0000c0; } .csharpcode .preproc { color: #cc6633; } .csharpcode .asp { background-color: #ffff00; } .csharpcode .html { color: #800000; } .csharpcode .attr { color: #ff0000; } .csharpcode .alt { background-color: #f4f4f4; width: 100%; margin: 0em; } .csharpcode .lnum { color: #606060; } .csharpcode, .csharpcode pre { font-size: small; color: black; font-family: consolas, "Courier New", courier, monospace; background-color: #ffffff; /*white-space: pre;*/ } .csharpcode pre { margin: 0em; } .csharpcode .rem { color: #008000; } .csharpcode .kwrd { color: #0000ff; } .csharpcode .str { color: #006080; } .csharpcode .op { color: #0000c0; } .csharpcode .preproc { color: #cc6633; } .csharpcode .asp { background-color: #ffff00; } .csharpcode .html { color: #800000; } .csharpcode .attr { color: #ff0000; } .csharpcode .alt { background-color: #f4f4f4; width: 100%; margin: 0em; } .csharpcode .lnum { color: #606060; }   3-Add the ModalPopupExtender ,moreover don’t forget to add the ScriptManager: <asp:ScriptManager ID="ScriptManager1" runat="server"/> <cc1:ModalPopupExtender ID="ModalPopupExtender1" runat="server" TargetControlID="btnView" PopupControlID="panModal" CancelControlID="btnCancel"/> .csharpcode, .csharpcode pre { font-size: small; color: black; font-family: consolas, "Courier New", courier, monospace; background-color: #ffffff; /*white-space: pre;*/ } .csharpcode pre { margin: 0em; } .csharpcode .rem { color: #008000; } .csharpcode .kwrd { color: #0000ff; } .csharpcode .str { color: #006080; } .csharpcode .op { color: #0000c0; } .csharpcode .preproc { color: #cc6633; } .csharpcode .asp { background-color: #ffff00; } .csharpcode .html { color: #800000; } .csharpcode .attr { color: #ff0000; } .csharpcode .alt { background-color: #f4f4f4; width: 100%; margin: 0em; } .csharpcode .lnum { color: #606060; }     4-In the pageLoad javascript function inside add_shown event set the focus on the txtName , and inside add_hiding reset the two Textboxes. <script language="javascript" type="text/javascript"> function pageLoad() { $find('ModalPopupExtender1').add_shown(function() { alert('add_shown event fires'); $get('<%=txtName.ClientID%>').focus();   });   $find('ModalPopupExtender1').add_hiding(function() { alert('add_hiding event fires'); $get('<%=txtName.ClientID%>').value = ""; $get('<%=txtPassword.ClientID%>').value = "";   }); }   </script> .csharpcode, .csharpcode pre { font-size: small; color: black; font-family: consolas, "Courier New", courier, monospace; background-color: #ffffff; /*white-space: pre;*/ } .csharpcode pre { margin: 0em; } .csharpcode .rem { color: #008000; } .csharpcode .kwrd { color: #0000ff; } .csharpcode .str { color: #006080; } .csharpcode .op { color: #0000c0; } .csharpcode .preproc { color: #cc6633; } .csharpcode .asp { background-color: #ffff00; } .csharpcode .html { color: #800000; } .csharpcode .attr { color: #ff0000; } .csharpcode .alt { background-color: #f4f4f4; width: 100%; margin: 0em; } .csharpcode .lnum { color: #606060; }   I’ve added the two alerts just to let you show when the event fires.   Hope this simple example show you the benefit and how to use these events.

  Read the article

• Pluralsight Meet the Author Podcast on Structuring JavaScript Code

  - by dwahlin

  I had the opportunity to talk with Fritz Onion from Pluralsight about one of my recent courses titled Structuring JavaScript Code for one of their Meet the Author podcasts. We talked about why JavaScript patterns are important for building more re-useable and maintainable apps, pros and cons of different patterns, and how to go about picking a pattern as a project is started. The course provides a solid walk-through of converting what I call “Function Spaghetti Code” into more modular code that’s easier to maintain, more re-useable, and less susceptible to naming conflicts. Patterns covered in the course include the Prototype Pattern, Revealing Module Pattern, and Revealing Prototype Pattern along with several other tips and techniques that can be used. Meet the Author:  Dan Wahlin on Structuring JavaScript Code   The transcript from the podcast is shown below: [Fritz]  Hello, this is Fritz Onion with another Pluralsight author interview. Today we’re talking with Dan Wahlin about his new course, Structuring JavaScript Code. Hi, Dan, it’s good to have you with us today. [Dan]  Thanks for having me, Fritz. [Fritz]  So, Dan, your new course, which came out in December of 2011 called Structuring JavaScript Code, goes into several patterns of usage in JavaScript as well as ways of organizing your code and what struck me about it was all the different techniques you described for encapsulating your code. I was wondering if you could give us just a little insight into what your motivation was for creating this course and sort of why you decided to write it and record it. [Dan]  Sure. So, I got started with JavaScript back in the mid 90s. In fact, back in the days when browsers that most people haven’t heard of were out and we had JavaScript but it wasn’t great. I was on a project in the late 90s that was heavy, heavy JavaScript and we pretty much did what I call in the course function spaghetti code where you just have function after function, there’s no rhyme or reason to how those functions are structured, they just kind of flow and it’s a little bit hard to do maintenance on it, you really don’t get a lot of reuse as far as from an object perspective. And so coming from an object-oriented background in JAVA and C#, I wanted to put something together that highlighted kind of the new way if you will of writing JavaScript because most people start out just writing functions and there’s nothing with that, it works, but it’s definitely not a real reusable solution. So the course is really all about how to move from just kind of function after function after function to the world of more encapsulated code and more reusable and hopefully better maintenance in the process. [Fritz]  So I am sure a lot of people have had similar experiences with their JavaScript code and will be looking forward to seeing what types of patterns you’ve put forth. Now, a couple I noticed in your course one is you start off with the prototype pattern. Do you want to describe sort of what problem that solves and how you go about using it within JavaScript? [Dan]  Sure. So, the patterns that are covered such as the prototype pattern and the revealing module pattern just as two examples, you know, show these kind of three things that I harp on throughout the course of encapsulation, better maintenance, reuse, those types of things. The prototype pattern specifically though has a couple kind of pros over some of the other patterns and that is the ability to extend your code without touching source code and what I mean by that is let’s say you’re writing a library that you know either other teammates or other people just out there on the Internet in general are going to be using. With the prototype pattern, you can actually write your code in such a way that we’re leveraging the JavaScript property and by doing that now you can extend my code that I wrote without touching my source code script or you can even override my code and perform some new functionality. Again, without touching my code.  And so you get kind of the benefit of the almost like inheritance or overriding in object oriented languages with this prototype pattern and it makes it kind of attractive that way definitely from a maintenance standpoint because, you know, you don’t want to modify a script I wrote because I might roll out version 2 and now you’d have to track where you change things and it gets a little tricky. So with this you just override those pieces or extend them and get that functionality and that’s kind of some of the benefits that that pattern offers out of the box. [Fritz]  And then the revealing module pattern, how does that differ from the prototype pattern and what problem does that solve differently? [Dan]  Yeah, so the prototype pattern and there’s another one that’s kind of really closely lined with revealing module pattern called the revealing prototype pattern and it also uses the prototype key word but it’s very similar to the one you just asked about the revealing module pattern. [Fritz]  Okay. [Dan]  This is a really popular one out there. In fact, we did a project for Microsoft that was very, very heavy JavaScript. It was an HMTL5 jQuery type app and we use this pattern for most of the structure if you will for the JavaScript code and what it does in a nutshell is allows you to get that encapsulation so you have really a single function wrapper that wraps all your other child functions but it gives you the ability to do public versus private members and this is kind of a sort of debate out there on the web. Some people feel that all JavaScript code should just be directly accessible and others kind of like to be able to hide their, truly their private stuff and a lot of people do that. You just put an underscore in front of your field or your variable name or your function name and that kind of is the defacto way to say hey, this is private. With the revealing module pattern you can do the equivalent of what objective oriented languages do and actually have private members that you literally can’t get to as an external consumer of the JavaScript code and then you can expose only those members that you want to be public. Now, you don’t get the benefit though of the prototype feature, which is I can’t easily extend the revealing module pattern type code if you don’t like something I’m doing, chances are you’re probably going to have to tweak my code to fix that because we’re not leveraging prototyping but in situations where you’re writing apps that are very specific to a given target app, you know, it’s not a library, it’s not going to be used in other apps all over the place, it’s a pattern I actually like a lot, it’s very simple to get going and then if you do like that public/private feature, it’s available to you. [Fritz]  Yeah, that’s interesting. So it’s almost, you can either go private by convention just by using a standard naming convention or you can actually enforce it by using the prototype pattern. [Dan]  Yeah, that’s exactly right. [Fritz]  So one of the things that I know I run across in JavaScript and I’m curious to get your take on is we do have all these different techniques of encapsulation and each one is really quite different when you’re using closures versus simply, you know, referencing member variables and adding them to your objects that the syntax changes with each pattern and the usage changes. So what would you recommend for people starting out in a brand new JavaScript project? Should they all sort of decide beforehand on what patterns they’re going to stick to or do you change it based on what part of the library you’re working on? I know that’s one of the points of confusion in this space. [Dan]  Yeah, it’s a great question. In fact, I just had a company ask me about that. So which one do I pick and, of course, there’s not one answer fits all. [Fritz]  Right. [Dan]  So it really depends what you just said is absolutely in my opinion correct, which is I think as a, especially if you’re on a team or even if you’re just an individual a team of one, you should go through and pick out which pattern for this particular project you think is best. Now if it were me, here’s kind of the way I think of it. If I were writing a let’s say base library that several web apps are going to use or even one, but I know that there’s going to be some pieces that I’m not really sure on right now as I’m writing I and I know people might want to hook in that and have some better extension points, then I would look at either the prototype pattern or the revealing prototype. Now, really just a real quick summation between the two the revealing prototype also gives you that public/private stuff like the revealing module pattern does whereas the prototype pattern does not but both of the prototype patterns do give you the benefit of that extension or that hook capability. So, if I were writing a library that I need people to override things or I’m not even sure what I need them to override, I want them to have that option, I’d probably pick a prototype, one of the prototype patterns. If I’m writing some code that is very unique to the app and it’s kind of a one off for this app which is what I think a lot of people are kind of in that mode as writing custom apps for customers, then my personal preference is the revealing module pattern you could always go with the module pattern as well which is very close but I think the revealing module patterns a little bit cleaner and we go through that in the course and explain kind of the syntax there and the differences. [Fritz]  Great, that makes a lot of sense. [Fritz]  I appreciate you taking the time, Dan, and I hope everyone takes a chance to look at your course and sort of make these decisions for themselves in their next JavaScript project. Dan’s course is, Structuring JavaScript Code and it’s available now in the Pluralsight Library. So, thank you very much, Dan. [Dan]  Thanks for having me again.

  Read the article

• Parallelism in .NET – Part 11, Divide and Conquer via Parallel.Invoke

  - by Reed

  Many algorithms are easily written to work via recursion.  For example, most data-oriented tasks where a tree of data must be processed are much more easily handled by starting at the root, and recursively “walking” the tree.  Some algorithms work this way on flat data structures, such as arrays, as well.  This is a form of divide and conquer: an algorithm design which is based around breaking up a set of work recursively, “dividing” the total work in each recursive step, and “conquering” the work when the remaining work is small enough to be solved easily. Recursive algorithms, especially ones based on a form of divide and conquer, are often a very good candidate for parallelization. This is apparent from a common sense standpoint.  Since we’re dividing up the total work in the algorithm, we have an obvious, built-in partitioning scheme.  Once partitioned, the data can be worked upon independently, so there is good, clean isolation of data. Implementing this type of algorithm is fairly simple.  The Parallel class in .NET 4 includes a method suited for this type of operation: Parallel.Invoke.  This method works by taking any number of delegates defined as an Action, and operating them all in parallel.  The method returns when every delegate has completed: Parallel.Invoke( () => { Console.WriteLine("Action 1 executing in thread {0}", Thread.CurrentThread.ManagedThreadId); }, () => { Console.WriteLine("Action 2 executing in thread {0}", Thread.CurrentThread.ManagedThreadId); }, () => { Console.WriteLine("Action 3 executing in thread {0}", Thread.CurrentThread.ManagedThreadId); } ); .csharpcode, .csharpcode pre { font-size: small; color: black; font-family: consolas, "Courier New", courier, monospace; background-color: #ffffff; /*white-space: pre;*/ } .csharpcode pre { margin: 0em; } .csharpcode .rem { color: #008000; } .csharpcode .kwrd { color: #0000ff; } .csharpcode .str { color: #006080; } .csharpcode .op { color: #0000c0; } .csharpcode .preproc { color: #cc6633; } .csharpcode .asp { background-color: #ffff00; } .csharpcode .html { color: #800000; } .csharpcode .attr { color: #ff0000; } .csharpcode .alt { background-color: #f4f4f4; width: 100%; margin: 0em; } .csharpcode .lnum { color: #606060; } Running this simple example demonstrates the ease of using this method.  For example, on my system, I get three separate thread IDs when running the above code.  By allowing any number of delegates to be executed directly, concurrently, the Parallel.Invoke method provides us an easy way to parallelize any algorithm based on divide and conquer.  We can divide our work in each step, and execute each task in parallel, recursively. For example, suppose we wanted to implement our own quicksort routine.  The quicksort algorithm can be designed based on divide and conquer.  In each iteration, we pick a pivot point, and use that to partition the total array.  We swap the elements around the pivot, then recursively sort the lists on each side of the pivot.  For example, let’s look at this simple, sequential implementation of quicksort: public static void QuickSort<T>(T[] array) where T : IComparable<T> { QuickSortInternal(array, 0, array.Length - 1); } private static void QuickSortInternal<T>(T[] array, int left, int right) where T : IComparable<T> { if (left >= right) { return; } SwapElements(array, left, (left + right) / 2); int last = left; for (int current = left + 1; current <= right; ++current) { if (array[current].CompareTo(array[left]) < 0) { ++last; SwapElements(array, last, current); } } SwapElements(array, left, last); QuickSortInternal(array, left, last - 1); QuickSortInternal(array, last + 1, right); } static void SwapElements<T>(T[] array, int i, int j) { T temp = array[i]; array[i] = array[j]; array[j] = temp; } Here, we implement the quicksort algorithm in a very common, divide and conquer approach.  Running this against the built-in Array.Sort routine shows that we get the exact same answers (although the framework’s sort routine is slightly faster).  On my system, for example, I can use framework’s sort to sort ten million random doubles in about 7.3s, and this implementation takes about 9.3s on average. Looking at this routine, though, there is a clear opportunity to parallelize.  At the end of QuickSortInternal, we recursively call into QuickSortInternal with each partition of the array after the pivot is chosen.  This can be rewritten to use Parallel.Invoke by simply changing it to: // Code above is unchanged... SwapElements(array, left, last); Parallel.Invoke( () => QuickSortInternal(array, left, last - 1), () => QuickSortInternal(array, last + 1, right) ); } This routine will now run in parallel.  When executing, we now see the CPU usage across all cores spike while it executes.  However, there is a significant problem here – by parallelizing this routine, we took it from an execution time of 9.3s to an execution time of approximately 14 seconds!  We’re using more resources as seen in the CPU usage, but the overall result is a dramatic slowdown in overall processing time. This occurs because parallelization adds overhead.  Each time we split this array, we spawn two new tasks to parallelize this algorithm!  This is far, far too many tasks for our cores to operate upon at a single time.  In effect, we’re “over-parallelizing” this routine.  This is a common problem when working with divide and conquer algorithms, and leads to an important observation: When parallelizing a recursive routine, take special care not to add more tasks than necessary to fully utilize your system. This can be done with a few different approaches, in this case.  Typically, the way to handle this is to stop parallelizing the routine at a certain point, and revert back to the serial approach.  Since the first few recursions will all still be parallelized, our “deeper” recursive tasks will be running in parallel, and can take full advantage of the machine.  This also dramatically reduces the overhead added by parallelizing, since we’re only adding overhead for the first few recursive calls.  There are two basic approaches we can take here.  The first approach would be to look at the total work size, and if it’s smaller than a specific threshold, revert to our serial implementation.  In this case, we could just check right-left, and if it’s under a threshold, call the methods directly instead of using Parallel.Invoke. The second approach is to track how “deep” in the “tree” we are currently at, and if we are below some number of levels, stop parallelizing.  This approach is a more general-purpose approach, since it works on routines which parse trees as well as routines working off of a single array, but may not work as well if a poor partitioning strategy is chosen or the tree is not balanced evenly. This can be written very easily.  If we pass a maxDepth parameter into our internal routine, we can restrict the amount of times we parallelize by changing the recursive call to: // Code above is unchanged... SwapElements(array, left, last); if (maxDepth < 1) { QuickSortInternal(array, left, last - 1, maxDepth); QuickSortInternal(array, last + 1, right, maxDepth); } else { --maxDepth; Parallel.Invoke( () => QuickSortInternal(array, left, last - 1, maxDepth), () => QuickSortInternal(array, last + 1, right, maxDepth)); } We no longer allow this to parallelize indefinitely – only to a specific depth, at which time we revert to a serial implementation.  By starting the routine with a maxDepth equal to Environment.ProcessorCount, we can restrict the total amount of parallel operations significantly, but still provide adequate work for each processing core. With this final change, my timings are much better.  On average, I get the following timings: Framework via Array.Sort: 7.3 seconds Serial Quicksort Implementation: 9.3 seconds Naive Parallel Implementation: 14 seconds Parallel Implementation Restricting Depth: 4.7 seconds Finally, we are now faster than the framework’s Array.Sort implementation.

  Read the article

• Turn Photos and Home Videos into Movies with Windows Live Movie Maker

  - by DigitalGeekery

  Are you looking for an easy way to take your digital photos and videos and turn them into a movie or slideshow? Today we’ll take a detailed look at how to do use Windows Live Movie Maker. Installation Windows Live Movie Maker comes bundled as part of the Windows Live Essentials suite (link below). However, you don’t have to install any of the programs you may not want. Take notice of the You’re almost done screen. Before clicking Continue, be sure to uncheck the boxes to set your search provider and homepage. Adding Pictures and Videos Open Windows Live Movie Maker. You can add videos or photos by simply dragging and dropping them onto the storyboard area. You can also click on the storyboard area or on the Add videos and photos button on the Home tab to browse for videos and photos. Windows Live Movie Maker supports most video, image, and audio file types. Select your files and add click Open to add them to Windows Live Movie Maker. By default WLMM doesn’t allow you to add files from network locations…so check out our article on how to add network support to Windows Live MovieMaker if the files you want to add are on a network drive. Layout All of your added clips will appear in the storyboard area on the right, while the currently selected clip will appear in the preview window on the left. You can adjust the size of the two areas by clicking and dragging the dividing line in the middle.    Make the clips on the storyboard bigger or smaller by clicking on the thumbnail size icon. The slider at the lower right adjusts the zoom time scale.   Previewing your Movie At any time, you can playback your movie and preview how it will look in the Preview window by clicking the space bar, or by pushing the play button under the preview window. You can also manually move the preview bar slider across the storyboard to view the clips as the video progresses. Adjusting Clips on the Storyboard You can click and drag clips on the storyboard to change the order in which the photos and videos appear.   Adding Music Nothing brings a movie to life quite like music. Selecting Add music will add your music to the beginning of the movie. Select Add music at the current point to include it in the movie to the current location of your preview bar slider, then browse for your music clip. WLMM supports many common audio files such as WAV, MP3, M4A, WMA, AIFF, and ASF. The music clip will appear above the video / photos clips on the storyboard.   You can change the location of music clips by clicking and dragging them to a different location on the storyboard. Add Titles, Captions, and Credits To add a Title screen to your movie, click the Title button on the Home tab. Type your title directly into the text box on the preview screen. The title will be placed at the location of the preview slider on the storyboard. However, you can change the location by clicking and dragging title to other areas of the storyboard. On the Format tab, there are a handful of text settings. You can change the font, color, size, alignment,  and transparency. The Adjust group allows you to change the background color, edit the text, and set the length of time the Title will appear in the movie.   The Effects group on the Format tab allows you to select an effect for your title screen. By hovering your cursor over each option, you will get a live preview of how each effect will appear in the preview window. Click to apply any of the effects. For captions, select where you want your caption to appear with the preview slider on the storyboard, then click the captions button on the Home tab. Just like the title, you type your caption directly into the text box on the preview screen, and you can make any adjustments by using the Font and Paragraph, Adjust, and Effects groups above. Credits are done the same as titles and captions, except they are automatically placed at the end of the movie.   Transitions Go to the Animation tab on the ribbon to apply transitions. Select a clip from the storyboard and hover over one of the transition to see it in the preview window. Click on the transition to apply it to the clip. You can apply transitions separately to clips or hold down Ctrl button while clicking to select multiple clips to which to apply the same transition. Pan and zoom effects are also located on the Animations tab, but can be applied to photos only. Like transition, you can apply them individually to a clip or hold down Ctrl button while clicking to select multiple clips to which to apply the same pan and zoom effect. Once applied, you can adjust the duration of the transitions and pan and zoom effects. You can also click the dropdown for additional transitions or effects. Visual Effects Similar to Pan and Zoom and Transitions, you can apply a variety of Visual Effects to individual or multiple clips. Editing Video and Music Note: This does not actually edit the original video you imported into your Windows Live Movie Maker project, only how it appears in your WLMM project. There are some very basic editing tools located on the Home tab. The Rotate left and Rotate right button will adjust any clip that may be oriented incorrectly. The Fit to music button will automatically adjust the duration of the photos (if you have any in your project) to fit the length of the music in your movie. Audio mix allows you to change the volume level   You can also do some slightly more advanced editing from the Edit tab. Select the video clip on the storyboard and click the Trim tool to edit or remove portions of a video clip. Next, click and drag the sliders in the preview windows to select the are you wish to keep. For example, the area outside the sliders is the area trimmed from the movie. The area inside is the section that is kept in the movie. You can also adjust the Start and End points manually on the ribbon.   When you are finished, click Save trim. You can also split your video clips. Move the preview slider to the location in the video clip where you’d like to split it, and select Split. Your video will be split into separate sections. Now you can apply different effects or move them to different locations on the storyboard. Editing Music Clips Select the music clip on the storyboard and then the Options tab on the ribbon. You can adjust the music volume by moving the slider right and left.   You can also choose to have your music clip fade in or out at the beginning and end of your movie. From the Fade in and Fade out dropdowns, select None, Slow, Medium, or Fast. To adjust the sound of your audio clips, click on the Edit tab, select the Video volume button, and adjust the slider. Move it all the way to the left to mute any background noise in your video clips.   AutoMovie As you have seen, Windows Live Movie Maker allows you to add effects, transitions, titles, and more. If you don’t want to do any of that stuff yourself, AutoMovie will automatically add title, credits, cross fade transitions between items, pan and zoom effects to photos, and fit your project to the music. Just select the AutoMovie button on the Home tab. You can go from zero to movie in literally a couple minutes.   Uploading to YouTube You can share your video on YouTube directly from Windows Live Movie Maker. Click on the YouTube icon in the Sharing group on the Home tab. You’ll be prompted for your YouTube username and password. Fill in the details about your movie and click Publish. The movie will be converted to WMV before being uploaded to YouTube. As soon as the YouTube conversion is complete, you’re new movie is live and ready to be viewed. Saving your Movie as a Video File Select the icon at the top left, then select Save movie. As you hover your mouse over each of the options, you will see the output display size, aspect ratio, and estimated file size per minute of video. All of these settings will output your movie as a WMV file. (Unfortunately, the only option is to save a movie as a WMV file.) The only difference is how they are encoded based on preset common settings. The Burn to DVD option also outputs a WMV file, but then opens Windows DVD Maker and walks you through the process of creating and burning a DVD.   If you choose the Burn to DVD option, close this window when the WMV file conversion is complete and the Windows DVD Maker will prompt you to begin. When your movie is finished, it’s time to relax and enjoy.   Conclusion Windows Live Movie Maker makes it easy for the average person to quickly churn out nice looking movies and slideshows from there own pictures and videos. However, long time users of previous editions (formerly called Windows Movie Maker) will likely be disappointed by some features missing in Windows Live Movie Maker that existed in earlier editions. Looking for details on burning your new project to DVD, check out our article on how to create and author DVDs with Windows DVD Maker. Download Windows Live Movie Maker Similar Articles Productive Geek Tips Family Fun: Share Photos with Photo Gallery and Windows Live SpacesCreate and Author DVDs in Windows 7Rotate a Video 90 degrees with VLC or Windows Live Movie MakerInstall Windows Live Essentials In Windows 7How to Make/Edit a movie with Windows Movie Maker in Windows Vista TouchFreeze Alternative in AutoHotkey The Icy Undertow Desktop Windows Home Server – Backup to LAN The Clear & Clean Desktop Use This Bookmarklet to Easily Get Albums Use AutoHotkey to Assign a Hotkey to a Specific Window Latest Software Reviews Tinyhacker Random Tips CloudBerry Online Backup 1.5 for Windows Home Server Snagit 10 VMware Workstation 7 Acronis Online Backup Windows Firewall with Advanced Security – How To Guides Sculptris 1.0, 3D Drawing app AceStock, a Tiny Desktop Quote Monitor Gmail Button Addon (Firefox) Hyperwords addon (Firefox) Backup Outlook 2010

  Read the article

• What approaches exist to setting up continent/country/city drop down menus?

  - by Dave

  How easy (or difficult) is it to have a Continent/Country/City drop down menu? Where one select from Drop Down Menus (for example): 1 - Europe 2 - UK 3 - London and then writes the Province/Area (for example: Essex). Realistically, how long should it take an experienced web developer to write the code of the above, as well as to link this selection to a Browse function and database storing? I do not have a geographical database yet and I am wondering what the fastest and cheapest way to add it to the drop down menu is. Is there any way to get that geographical database for free? I can see this type of geographical drop down menu in thousands of websites, but I am struggling as to how to implement it ASAP. Follow Up: Tks All x your answers and comments so far. I hear what you are saying. I understand that there are rare occasions of Countries with multiple (same) name Cities and that it might be disputable whether a Country belongs to a certain Continent/Region or not (see Russia x example, Europe or Asia?). Anyway, please take a look, for instance, at this website Sign UP screen http://www.couchsurfing.org/register.html My question then is: Where do I get that list (Country/Cities) and how do I create that _array? Manually copying it somewhere else (which would take me ages) or are there ready made lists that can be downloaded from somewhere for free?

  Read the article

• The HTG Guide to Using a Bluetooth Keyboard with Your Android Device

  - by Matt Klein

  Android devices aren’t usually associated with physical keyboards. But, since Google is now bundling their QuickOffice app with the newly-released Kit-Kat, it appears inevitable that at least some Android tablets (particularly 10-inch models) will take on more productivity roles. In recent years, physical keyboards have been rendered obsolete by swipe style input methods such as Swype and Google Keyboard. Physical keyboards tend to make phones thick and plump, and that won’t fly today when thin (and even flexible and curved) is in vogue. So, you’ll be hard-pressed to find smartphone manufacturers launching new models with physical keyboards, thus rendering sliders to a past chapter in mobile phone evolution. It makes sense to ditch the clunky keyboard phone in favor of a lighter, thinner model. You’re going to carry around in your pocket or purse all day, why have that extra bulk and weight? That said, there is sound logic behind pairing tablets with keyboards. Microsoft continues to plod forward with its Surface models, and while critics continue to lavish praise on the iPad, its functionality is obviously enhanced and extended when you add a physical keyboard. Apple even has an entire page devoted specifically to iPad-compatible keyboards. But an Android tablet and a keyboard? Does such a thing even exist? They do actually. There are docking keyboards and keyboard/case combinations, there’s the Asus Transformer family, Logitech markets a Windows 8 keyboard that speaks “Android”, and these are just to name a few. So we know that keyboard products that are designed to work with Android exist, but what about an everyday Bluetooth keyboard you might use with Windows or OS X? How-To Geek wanted look at how viable it is to use such a keyboard with Android. We conducted some research and examined some lists of Android keyboard shortcuts. Most of what we found was long outdated. Many of the shortcuts don’t even apply anymore, while others just didn’t work. Regardless, after a little experimentation and a dash of customization, it turns out using a keyboard with Android is kind of fun, and who knows, maybe it will catch on. Setting things up Setting up a Bluetooth keyboard with Android is very easy. First, you’ll need a Bluetooth keyboard and of course an Android device, preferably running version 4.1 (Jelly Bean) or higher. For our test, we paired a second-generation Google Nexus 7 running Android 4.3 with a Samsung Series 7 keyboard. In Android, enable Bluetooth if it isn’t already on. We’d like to note that if you don’t normally use Bluetooth accessories and peripherals with your Android device (or any device really), it’s best practice to leave Bluetooth off because, like GPS, it drains the device’s battery more quickly. To enable Bluetooth, simply go to “Settings” -> “Bluetooth” and tap the slider button to “On”. To set up the keyboard, make sure it is on and then tap “Bluetooth” in the Android settings. On the resulting screen, your Android device should automatically search for and hopefully find your keyboard. If you don’t get it right the first time, simply turn the keyboard on again and then tap “Search for Devices” to try again. If it still doesn’t work, make sure you have fresh batteries and the keyboard isn’t paired to another device. If it is, you will need to unpair it before it will work with your Android device (consult your keyboard manufacturer’s documentation or Google if you don’t know how to do this). When Android finds your keyboard, select it under “Available Devices” … … and you should be prompted to type in a code: If successful, you will see that device is now “Connected” and you’re ready to go. If you want to test things out, try pressing the “Windows” key (“Apple” or “Command”) + ESC, and you will be whisked to your Home screen. So, what can you do? Traditional Mac and Windows users know there’s usually a keyboard shortcut for just about everything (and if there isn’t, there’s all kinds of ways to remap keys to do a variety of commands, tasks, and functions). So where does Android fall in terms of baked-in keyboard commands? There answer to that is kind of enough, but not too much. There are definitely established combos you can use to get around, but they aren’t clear and there doesn’t appear to be any one authority on what they are. Still, there is enough keyboard functionality in Android to make it a viable option, if only for those times when you need to get something done (long e-mail or important document) and an on-screen keyboard simply won’t do. It’s important to remember that Android is, and likely always will be a touch-first interface. That said, it does make some concessions to physical keyboards. In other words, you can get around Android fairly well without having to lift your hands off the keys, but you will still have to tap the screen regularly, unless you add a mouse. For example, you can wake your device by tapping a key rather than pressing its power button. However, if your device is slide or pattern-locked, then you’ll have to use the touchscreen to unlock it – a password or PIN however, works seamlessly with a keyboard – other things like widgets and app controls and features, have to be tapped. You get the idea. Keyboard shortcuts and navigation As we said, baked-in keyboard shortcut combos aren’t necessarily abundant nor apparent. The one thing you can always do is search. Any time you want to Google something, start typing from the Home screen and the search screen will automatically open and begin displaying results. Other than that, here is what we were able to figure out: ESC = go back CTRL + ESC = menu CTRL + ALT + DEL = restart (no questions asked) ALT + SPACE = search page (say “OK Google” to voice search) ALT + TAB (ALT + SHIFT + TAB) = switch tasks Also, if you have designated volume function keys, those will probably work too. There’s also some dedicated app shortcuts like calculator, Gmail, and a few others: CMD + A = calculator CMD + C = contacts CMD + E = e-mail CMD + G = Gmail CMD + L = Calendar CMD + P = Play Music CMD + Y = YouTube Overall, it’s not a long comprehensive list and there’s no dedicated keyboard combos for the full array of Google’s products. Granted, it’s hard to imagine getting a lot of mileage out of a keyboard with Maps but with something like Keep, you could type out long, detailed lists on your tablet, and then view them on your smartphone when you go out shopping. You can also use the arrow keys to navigate your Home screen over shortcuts and open the app drawer. When something on the screen is selected, it will be highlighted in blue. Press “Enter” to open your selection. Additionally, if an app has its own set of shortcuts, e.g. Gmail has quite a few unique shortcuts to it, as does Chrome, some – though not many – will work in Android (not for YouTube though). Also, many “universal” shortcuts such as Copy (CTRL + C), Cut (CTRL + X), Paste (CTRL + V), and Select All (CTRL + A) work where needed – such as in instant messaging, e-mail, social media apps, etc. Creating custom application shortcuts What about custom shortcuts? When we were researching this article, we were under the impression that it was possible to assign keyboard combinations to specific apps, such as you could do on older Android versions such as Gingerbread. This no long seems to be the case and nowhere in “Settings” could we find a way to assign hotkey combos to any of our favorite, oft-used apps or functions. If you do want custom keyboard shortcuts, what can you do? Luckily, there’s an app on Google Play that allows you to, among other things, create custom app shortcuts. It is called External Keyboard Helper (EKH) and while there is a free demo version, the pay version is only a few bucks. We decided to give EKH a whirl and through a little experimentation and finally reading the developer’s how-to, we found we could map custom keyboard combos to just about anything. To do this, first open the application and you’ll see the main app screen. Don’t worry about choosing a custom layout or anything like that, you want to go straight to the “Advanced settings”: In the “Advanced settings” select “Application shortcuts” to continue: You can have up to 16 custom application shortcuts. We are going to create a custom shortcut to the Facebook app. We choose “A0”, and from the resulting list, Facebook. You can do this for any number of apps, services, and settings. As you can now see, the Facebook app has now been linked to application-zero (A0): Go back to the “Advanced settings” and choose “Customize keyboard mappings”: You will be prompted to create a custom keyboard layout so we choose “Custom 1”: When you choose to create a custom layout, you can do a great many more things with your keyboard. For example, many keyboards have predefined function (Fn) keys, which you can map to your tablet’s brightness controls, toggle WiFi on/off, and much more. A word of advice, the application automatically remaps certain keys when you create a custom layout. This might mess up some existing keyboard combos. If you simply want to add some functionality to your keyboard, you can go ahead and delete EKH’s default changes and start your custom layout from scratch. To create a new combo, select “Add new key mapping”: For our new shortcut, we are going to assign the Facebook app to open when we key in “ALT + F”. To do this, we press the “F” key while in the “Scancode” field and we see it returns a value of “33”. If we wanted to use a different key, we can press “Change” and scan another key’s numerical value. We now want to assign the “ALT” key to application “A0”, previously designated as the Facebook app. In the “AltGr” field, we enter “A0” and then “Save” our custom combo. And now we see our new application shortcut. Now, as long as we’re using our custom layout, every time we press “ALT + F”, the Facebook app will launch: External Keyboard Helper extends far beyond simple application shortcuts and if you are looking for deeper keyboard customization options, you should definitely check it out. Among other things, EKH also supports dozens of languages, allows you to quickly switch between layouts using a key or combo, add up to 16 custom text shortcuts, and much more! It can be had on Google Play for $2.53 for the full version, but you can try the demo version for free. More extensive documentation on how to use the app is also available. Android? Keyboard? Sure, why not? Unlike traditional desktop operating systems, you don’t need a physical keyboard and mouse to use a mobile operating system. You can buy an iPad or Nexus 10 or Galaxy Note, and never need another accessory or peripheral – they work as intended right out of the box. It’s even possible you can write the next great American novel on one these devices, though that might require a lot of practice and patience. That said, using a keyboard with Android is kind of fun. It’s not revelatory but it does elevate the experience. You don’t even need to add customizations (though they are nice) because there are enough existing keyboard shortcuts in Android to make it usable. Plus, when it comes to inputting text such as in an editor or terminal application, we fully advocate big, physical keyboards. Bottom line, if you’re looking for a way to enhance your Android tablet, give a keyboard a chance. Do you use your Android device for productivity? Is a physical keyboard an important part of your setup? Do you have any shortcuts that we missed? Sound off in the comments and let us know what you think.     

  Read the article

• The C++ Standard Template Library as a BDB Database (part 1)

  - by Gregory Burd

  If you've used C++ you undoubtedly have used the Standard Template Libraries. Designed for in-memory management of data and collections of data this is a core aspect of all C++ programs. Berkeley DB is a database library with a variety of APIs designed to ease development, one of those APIs extends and makes use of the STL for persistent, transactional data storage. dbstl is an STL standard compatible API for Berkeley DB. You can make use of Berkeley DB via this API as if you are using C++ STL classes, and still make full use of Berkeley DB features. Being an STL library backed by a database, there are some important and useful features that dbstl can provide, while the C++ STL library can't. The following are a few typical use cases to use the dbstl extensions to the C++ STL for data storage. When data exceeds available physical memory.Berkeley DB dbstl can vastly improve performance when managing a dataset which is larger than available memory. Performance suffers when the data can't reside in memory because the OS is forced to use virtual memory and swap pages of memory to disk. Switching to BDB's dbstl improves performance while allowing you to keep using STL containers. When you need concurrent access to C++ STL containers.Few existing C++ STL implementations support concurrent access (create/read/update/delete) within a container, at best you'll find support for accessing different containers of the same type concurrently. With the Berkeley DB dbstl implementation you can concurrently access your data from multiple threads or processes with confidence in the outcome. When your objects are your database.You want to have object persistence in your application, and store objects in a database, and use the objects across different runs of your application without having to translate them to/from SQL. The dbstl is capable of storing complicated objects, even those not located on a continous chunk of memory space, directly to disk without any unnecessary overhead. These are a few reasons why you should consider using Berkeley DB's C++ STL support for your embedded database application. In the next few blog posts I'll show you a few examples of this approach, it's easy to use and easy to learn.

  Read the article

• Unable to install Google Chrome

  - by Jordan

  I receive this in my terminal when I try to install using wget https://dl.google.com/linux/direct/google-chrome-stable_current_i386.deb sudo dpkg -i google-chrome* or sudo dpkg --install /Path/to/chrome.deb I receive Selecting previously unselected package google-chrome-stable. (Reading database ... 146911 files and directories currently installed.) Unpacking google-chrome-stable (from google-chrome-stable_current_i386.deb) ... dpkg: dependency problems prevent configuration of google-chrome-stable: google-chrome-stable depends on xdg-utils (>= 1.0.2). dpkg: error processing google-chrome-stable (--install): dependency problems - leaving unconfigured Processing triggers for man-db ... Processing triggers for bamfdaemon ... Rebuilding /usr/share/applications/bamf.index... Processing triggers for desktop-file-utils ... Processing triggers for gnome-menus ... Errors were encountered while processing: google-chrome-stable I then type sudo apt-get install -f And retry installation though it still does not install and I receive the same errors. I have also tried using: sudo apt-get install libxss1 libnspr4-0d libcurl3 Though the above doesn't work either.

  Read the article

• SQL SERVER – Index Created on View not Used Often – Limitation of the View 12

  - by pinaldave

  I have previously written on the subject SQL SERVER – The Limitations of the Views – Eleven and more…. This was indeed a very popular series and I had received lots of feedback on that topic. Today we are going to discuss something very interesting as well. During my recent performance tuning seminar in Hyderabad, I presented on the subject of Views. During the seminar, one of the attendees asked a question: We create a table and create a View on the top of it. On the same view, if we create Index, when querying View, will that index be used? The answer is NOT Always! (There is only one specific condition when it will be used. We will write about that later in the next post). Let us see the test case for the same. In our script we will do following: USE tempdb GO IF EXISTS (SELECT * FROM sys.views WHERE OBJECT_ID = OBJECT_ID(N'[dbo].[SampleView]')) DROP VIEW [dbo].[SampleView] GO IF EXISTS (SELECT * FROM sys.objects WHERE OBJECT_ID = OBJECT_ID(N'[dbo].[mySampleTable]') AND TYPE IN (N'U')) DROP TABLE [dbo].[mySampleTable] GO -- Create SampleTable CREATE TABLE mySampleTable (ID1 INT, ID2 INT, SomeData VARCHAR(100)) INSERT INTO mySampleTable (ID1,ID2,SomeData) SELECT TOP 100000 ROW_NUMBER() OVER (ORDER BY o1.name), ROW_NUMBER() OVER (ORDER BY o2.name), o2.name FROM sys.all_objects o1 CROSS JOIN sys.all_objects o2 GO -- Create View CREATE VIEW SampleView WITH SCHEMABINDING AS SELECT ID1,ID2,SomeData FROM dbo.mySampleTable GO -- Create Index on View CREATE UNIQUE CLUSTERED INDEX [IX_ViewSample] ON [dbo].[SampleView] ( ID2 ASC ) GO -- Select from view SELECT ID1,ID2,SomeData FROM SampleView GO Let us check the execution plan for the last SELECT statement. You can see from the execution plan. That even though we are querying View and the View has index, it is not really using that index. In the next post, we will see the significance of this View and where it can be helpful. Meanwhile, I encourage you to read my View series: SQL SERVER – The Limitations of the Views – Eleven and more…. Reference: Pinal Dave (http://blog.sqlauthority.com) Filed under: Pinal Dave, SQL, SQL Authority, SQL Query, SQL Scripts, SQL Server, SQL Tips and Tricks, SQL Training, SQL View, T SQL, Technology

  Read the article

• Fix Problems Upgrading Office 2010 Beta to RTM (Final) Release

  - by Mysticgeek

  There are several scenarios where you may run into trouble uninstalling the 2010 Beta and trying to install the RTM (final) release. Today we’ll cover the problems we ran into, and how to fix them. You would think upgrading from the Office 2010 Beta to the final release would be an easy process. Unfortunately, it’s not always that simple. In fact, we ran into three different scenarios where the install wasn’t smooth whatsoever. If you currently have the 2010 Beta installed, you have to remove it before you can install the RTM.  Here we’ll take a look at three different troublesome install scenarios we ran into, and how we fixed each one. Important Note: Before proceeding with any of these steps, make sure and backup your Outlook .pst files! Scenario 1 – Uninstall Office 2010 Beta & Fix Install Errors In this first scenario we have Office Professional Plus 2010 Beta 32-bit installed on a Windows 7 Home Premium 32-bit system. First try to uninstall the Office 2010 Beta by going into Control Panel and selecting Programs and Features. Scroll down to Microsoft Office Professional Plus 2010, right-click it and select Uninstall. Click Yes when the confirmation dialog box comes up. Wait while Office 2010 Beta uninstalls…the amount of time it takes will vary from system to system. To complete the uninstall process, a reboot is required. Fixing Setup Errors The problem is when you start the installation of the 2010 RTM… You get the following setup error even though you uninstalled the 2010 Beta. The problem is there are leftover Office apps or stand alone Office products. So, we need a utility that will clean them up for us.   Windows Installer Clean Up Utility Download and install the Clean Up Utility (link Below) following the defaults. After it’s installed you’ll find it in Start \ All Programs \ Windows Install Clean Up …go ahead and launch the utility. Now go through and remove all Office Programs or addins that you find in the list. Make sure you are just deleting Office apps and not something you need like Java for example. If you’re not sure what something is, doing a quick Google search should help you out. For instance we had the Office labs Ribbon Hero installed… just highlight and click Remove. Remove anything that has something to do with Office…then reboot your machine. Now, you should be able to begin the installation of Office 2010 RTM (Final) Release without any errors. If you do get an error during the install process, like this one telling us we have old version of Groove Server… Navigate to C:\Users\username\AppData\Local\Microsoft (where username is the computer name) and delete any existing MS Office folders. Then try the install again, this solved the problem in our first scenario. Scenario 2 – Not Being Able to Uninstall 2010 Beta from Programs and Features In this next scenario we have Office Professional Plus 2010 Beta 32-bit installed on a Windows 7 Home Premium 32-bit system. Another problem we ran into is not being able to uninstall the 2010 Beta from Programs and Features. When you go in to uninstall it, nothing happens. If you run into this problem, we again need to download and install the Windows Installer Clean Up Utility (link below) and manually uninstall the Beta. When you launch it, scroll down to Microsoft Office Professional Plus 2010 (Beta), highlight it and click Remove.   Click OK to the Warning Dialog box… If you see any other Office 2010, 2007, or 2003 entries you can hold the “Shift” key and highlight them all…then click Remove and click OK to the warning dialog. Now we need to delete some Registry settings. Click on Start and type regedit into the Search box and hit Enter. Navigate to HKEY_CURRENT_USER \ Software \ Microsoft \ Office and delete the folder. Then navigate to HKEY_LOCAL_MACHINE \ Software \ Microsoft \ Office and delete those keys as well. Now go into C:\Program Files and find any of these three folders…Microsoft Office, OfficeUpdate, or OfficeUpdate14…you might find one, two or all three. Either way just rename the folders with “_OLD” (without quotes) at the end. Then go into C:\Users\username\AppData\Local\Microsoft and delete any existing MS Office folders. Where in this example we have office, Office Labs, One Note…etc. Now we want to delete the contents of the Temp folder. Click on Start and type %temp% into the Search box and hit Enter. Use the key combination “Ctrl+A” to select all the files in this folder, then right-click and click Delete, or simply hit the Delete key. If you have some files that won’t delete, just skip them as they shouldn’t affect the Office install. Then empty the Recycle Bin and restart your machine. When you get back from the restart launch the Office 2010 RTM installer and you should be good to go with installation. Because we uninstalled the Office 2010 Beta manually, you may have some lingering blank icons that you’ll need to clean up. Scenario –3 Uninstall 2007 and Install 2010 32-Bit on x64 Windows 7 For this final scenario we are uninstalling Office Professional 2007 and installing Office Professional Plus 2010 32-Bit edition on a Windows Ultimate 64-bit computer. This machine actually had Office 2010 Beta 64-bit installed at one point also, it’s since been removed, and 2007 was reinstalled.  Go into Programs and Settings and uninstall Microsoft Office Professional 2007. Click Yes to the dialog box asking if you’re sure you want to uninstall it… Then wait while Office 2007 is uninstalled. The amount of time it takes will vary between systems. A restart is required to complete the process… Again we need to call upon the Windows Installer Clean Up Utility. Go through and delete any left over Office 2007 and 2010 entries. Click OK to the warning dialog that comes up. After that’s complete, navigate to HKEY_CURRENT_USER \ Software \ Microsoft \ Office and delete the folder. Then navigate to HKEY_LOCAL_MACHINE \ Software \ Microsoft \ Office and delete those keys as well. We still need to go into C:\Users\ username\AppData\ Local\ Microsoft (where username is the computer name) and delete any Office folders. In this example we have Outlook Connector, Office, and Outlook to delete. Now let’s delete the contents of the Temp folder by typing %temp% into the Search box in the Start Menu. Then delete all of the files and folders in the Temp directory. If you have some files that won’t delete, just skip them as they shouldn’t affect the Office install. Then empty the Recycle Bin and restart your machine. If you try to install the 2010 RTM at this point you might be able to begin the install, but may get the following Error 1402 message. To solve this issue, we opened the command prompt and ran the following: secedit /configure /cfg %windir%\inf\defltbase.inf /db defltbase.sdb /verbose After the command completes, kick off the Office 2010 (Final) RTM 32-bit edition. This solved the issue and Office 2010 installed successfully.   Conclusion Except for the final scenario, we found using the Windows Installer Clean Up Utility to come in very handy. Using that along with deleting a couple folders and registry settings did the trick. In the last one, we had to get a bit more geeky and use some command line magic, but it got the job done. After some extensive testing in our labs, the only time the upgrade to the RTM went smoothly was when we had a clean Vista or Windows 7 system with a fresh install of the 2010 beta only. However, chances are you went from 2003 or 2007 to the free 2010 Beta. You might also have addins or other Office products installed, so there are going to be a lot of different office files scattered throughout your PC. If that’s the case, you may run into the issues we covered here. These are a few scenarios where we got errors and were not able to install Office 2010 after removing the beta. There could be other problems, and if any of you have experienced different issues or have more good suggestions, leave a comment and let us know! Link Download Windows Installer Clean Up Utility Similar Articles Productive Geek Tips Remove Office 2010 Beta and Reinstall Office 2007How to Upgrade the Windows 7 RC to RTM (Final Release)Upgrading Ubuntu from Dapper to Edgy with Update ManagerDisable Office 2010 Beta Send-a-Smile from StartupAdd or Remove Apps from the Microsoft Office 2007 or 2010 Suite TouchFreeze Alternative in AutoHotkey The Icy Undertow Desktop Windows Home Server – Backup to LAN The Clear & Clean Desktop Use This Bookmarklet to Easily Get Albums Use AutoHotkey to Assign a Hotkey to a Specific Window Latest Software Reviews Tinyhacker Random Tips DVDFab 6 Revo Uninstaller Pro Registry Mechanic 9 for Windows PC Tools Internet Security Suite 2010 Case Study – How to Optimize Popular Wordpress Sites Restore Hidden Updates in Windows 7 & Vista Iceland an Insurance Job? Find Downloads and Add-ins for Outlook Recycle ! Find That Elusive Icon with FindIcons

  Read the article

• SQL SERVER – Difference Between DATETIME and DATETIME2 – WITH GETDATE

  - by pinaldave

  Earlier I wrote blog post SQL SERVER – Difference Between GETDATE and SYSDATETIME which inspired me to write SQL SERVER – Difference Between DATETIME and DATETIME2. Now earlier two blog post inspired me to write this blog post (and 4 emails and 3 reads from readers). I previously populated DATETIME and DATETIME2 field with SYSDATETIME, which gave me very different behavior as SYSDATETIME was rounded up/down for the DATETIME datatype. I just ran the same experiment but instead of populating SYSDATETIME in this script I will be using GETDATE function. DECLARE @Intveral INT SET @Intveral = 10000 CREATE TABLE #TimeTable (FirstDate DATETIME, LastDate DATETIME2) WHILE (@Intveral > 0) BEGIN INSERT #TimeTable (FirstDate, LastDate) VALUES (GETDATE(), GETDATE()) SET @Intveral = @Intveral - 1 END GO SELECT COUNT(DISTINCT FirstDate) D_FirstDate, COUNT(DISTINCT LastDate) D_LastDate FROM #TimeTable GO SELECT DISTINCT a.FirstDate, b.LastDate FROM #TimeTable a INNER JOIN #TimeTable b ON a.FirstDate = b.LastDate GO SELECT * FROM #TimeTable GO DROP TABLE #TimeTable GO Let us run above script and observe the results. You will find that the values of GETDATE which is populated in both the columns FirstDate and LastDate are very much same. This is because GETDATE is of datatype DATETIME and the precision of the GETDATE is smaller than DATETIME2 there is no rounding happening. In other word, this experiment is pointless. I have included this as I got 4 emails and 3 twitter questions on this subject. If your datatype of variable is smaller than column datatype there is no manipulation of data, if data type of variable is larger than column datatype the data is rounded. Reference: Pinal Dave (http://www.SQLAuthority.com) Filed under: Pinal Dave, SQL, SQL Authority, SQL DateTime, SQL Query, SQL Scripts, SQL Server, SQL Tips and Tricks, T SQL, Technology

  Read the article

• Shared Development Space

  - by PatrickWalker

  Currently the company I work in gives each developer their own development virtual machine. On this machine (Windows 7) they install the entire stack of the product (minus database) this stack is normally spread amongst multiple machines of differing OS (although moving towards windows 2008 and 2008r2) So when a developer has a new project they are likely to be updating only a small piece of their stack and as such the rest of it can become out of date with the latest production code. The isolation from others means some issues won't be found until the code goes into shared test environments/production. I'm suggesting a move from functional testing on these isolated machines to plugging machines into a shared environment. The goal being to move towards a deployment thats closer to production in mechanism and server type. Developers would still make code changes on their Win7 vm and run unit/component testing locally but for functionally testing they would leverage a shared enviornment. Does anyone else use a shared development environment like this? Are there many reasons against this sort of sandbox environment? The biggest drawback is a move away from only checking in code when you've done local functional testing to checking in after static testing. I'm hoping an intelligent git branching strategy can take care of this for us.

  Read the article

• Envista: Coordinating Utilities with Oracle Spatial 11g

  - by stephen.garth

  It's annoying when the same streets seem to be perpetually dug up for utility construction or maintenance by your water or sewer department, electric utility, gas company or telephone company. Can't they do a better job of coordinating these activities? In this podcast, Marc Fagan, Executive VP of Product Management from Envista describes a Software-as-a-Service solution that Envista provides for utilities and public works departments to coordinate upcoming construction work, using Oracle Database 11g with Oracle Spatial. Each participating utility enters key data into the Web-based application, including when and where their work is to take place, and who to contact for more information. The data is then available on a common base map, enabling all participants to coordinate their activities, save money, and minimize inconvenience to their customers. Listen to the podcast Find out more about Oracle Spatial 11g var gaJsHost = (("https:" == document.location.protocol) ? "https://ssl." : "http://www."); document.write(unescape("%3Cscript src='" + gaJsHost + "google-analytics.com/ga.js' type='text/javascript'%3E%3C/script%3E")); try { var pageTracker = _gat._getTracker("UA-13185312-1"); pageTracker._trackPageview(); } catch(err) {}

  Read the article

• LLBLGen Pro v3.0 has been released!

  - by FransBouma

  After two years of hard work we released v3.0 of LLBLGen Pro today! V3.0 comes with a completely new designer which has been developed from the ground up for .NET 3.5 and higher. Below I'll briefly mention some highlights of this new release: Entity Framework (v1 & v4) support NHibernate support (hbm.xml mappings & FluentNHibernate mappings) Linq to SQL support Allows both Model first and Database first development, or a mixture of both .NET 4.0 support Model views Grouping of project elements Linq-based project search Value Type (DDD) support Multiple Database types in single project XML based project file Integrated template editor Relational Model Data management Flexible attribute declaration for code generation, no more buddy classes needed Fine-grained project validation Update / Create DDL SQL scripts Fast Text-DSL based Quick mode Powerful text-DSL based Quick Model functionality Per target framework extensible settings framework much much more... Of course we still support our own O/R mapper framework: LLBLGen Pro v3.0 Runtime framework as well, which was updated with some minor features and was upgraded to use the DbProviderFactory system. Please watch the videos of the designer (more to come very soon!) to see some aspects of the new designer in action. The full version comes with Algorithmia in sourcecode as well. Algorithmia is an algorithm library written for .NET 3.5 which powers the heart of the designer with a fine-grained undo/redo command framework, graph classes and much more. I'd like to thank all beta-testers, our support team and others who have helped us with this massive release. :)

  Read the article

• Step by Step Install of MAAS and JUJU

  - by John S

  I am working on understanding the pieces that I am missing in being able to deploy Juju across the other MAAS nodes. I don't know If I have a step out of place, or missing a few. The server owns the router which handles the DHCP and DNS. Any assistance is greatly appreciated. When I am at the end I will either get a 409 error, or arbitrary pick tools 1.16.0 error. It is worth mentioning that local, and aws works fine. Hopefully with all of these steps spelled out it will help someone else along the way too. Steps Setting Up MAAS and JUJU - 12.04 LTS Clean install SSH only from the package selection during install sudo apt-get install software-properties-common sudo apt-get install python-software-properties sudo add-apt-repository ppa:maas-maintainers/stable sudo add-apt-repository ppa:juju/stable sudo apt-get update sudo apt-get dist-upgrade sudo reboot sudo apt-get install maas maas-dns maas-dhcp sudo ufw disable sudo reboot - edit /etc/dhcp/dhcpd.conf authoritive subnet 10.0.0.0 netmask 255.255.255.0 { next-server 10.0.0.2; filename "pxelinux.0"; } sudo maas createsuperuser sudo maas-import-pxe-files Login to MAAS http://10.x.x.x/MAAS cluster controller configuration for eth0 manage dhcp and dns IP 10.0.0.2 subnet 255.255.255.0 broadcast 10.0.0.0 routerip 10.0.0.1 ip low 10.0.0.5 ip high 10.0.0.180 Commissioning default and distro is set at 12.04 default domain is at local sudo maas-cli login maas http://10.x.x.x/MAAS/api/1.0 api-key ssh-keygen -t rsa -b 2048 - enter - no password - cat id_rsa.pub and enter key into MAAS ssh sudo maas-cli maas nodes accept-all (interestingly enough I only get back [] when executing this ) PXE one machine, accept and commision, start and deploy. sudo apt-get install juju-core juju-local MAAS config: maas: type: maas maas-server: '://10.x.x.x:80/MAAS' maas-oauth: 'MAAS_API_KEY' admin-secret: 'nothing' default-series: 'precise' juju switch maas sudo juju bootstrap --show-log

  Read the article

• Back to Basics: When does a .NET Assembly Dependency get loaded

  - by Rick Strahl

  When we work on typical day to day applications, it's easy to forget some of the core features of the .NET framework. For me personally it's been a long time since I've learned about some of the underlying CLR system level services even though I rely on them on a daily basis. I often think only about high level application constructs and/or high level framework functionality, but the low level stuff is often just taken for granted. Over the last week at DevConnections I had all sorts of low level discussions with other developers about the inner workings of this or that technology (especially in light of my Low Level ASP.NET Architecture talk and the Razor Hosting talk). One topic that came up a couple of times and ended up a point of confusion even amongst some seasoned developers (including some folks from Microsoft <snicker>) is when assemblies actually load into a .NET process. There are a number of different ways that assemblies are loaded in .NET. When you create a typical project assemblies usually come from: The Assembly reference list of the top level 'executable' project The Assembly references of referenced projects Dynamically loaded at runtime via AppDomain/Reflection loading In addition .NET automatically loads mscorlib (most of the System namespace) the boot process that hosts the .NET runtime in EXE apps, or some other kind of runtime hosting environment (runtime hosting in servers like IIS, SQL Server or COM Interop). In hosting environments the runtime host may also pre-load a bunch of assemblies on its own (for example the ASP.NET host requires all sorts of assemblies just to run itself, before ever routing into your user specific code). Assembly Loading The most obvious source of loaded assemblies is the top level application's assembly reference list. You can add assembly references to a top level application and those assembly references are then available to the application. In a nutshell, referenced assemblies are not immediately loaded - they are loaded on the fly as needed. So regardless of whether you have an assembly reference in a top level project, or a dependent assembly assemblies typically load on an as needed basis, unless explicitly loaded by user code. The same is true of dependent assemblies. To check this out I ran a simple test: I have a utility assembly Westwind.Utilities which is a general purpose library that can work in any type of project. Due to a couple of small requirements for encoding and a logging piece that allows logging Web content (dependency on HttpContext.Current) this utility library has a dependency on System.Web. Now System.Web is a pretty large assembly and generally you'd want to avoid adding it to a non-Web project if it can be helped. So I created a Console Application that loads my utility library: You can see that the top level Console app a reference to Westwind.Utilities and System.Data (beyond the core .NET libs). The Westwind.Utilities project on the other hand has quite a few dependencies including System.Web. I then add a main program that accesses only a simple utillity method in the Westwind.Utilities library that doesn't require any of the classes that access System.Web: static void Main(string[] args) { Console.WriteLine(StringUtils.NewStringId()); Console.ReadLine(); } StringUtils.NewStringId() calls into Westwind.Utilities, but it doesn't rely on System.Web. Any guesses what the assembly list looks like when I stop the code on the ReadLine() command? I'll wait here while you think about it… … … So, when I stop on ReadLine() and then fire up Process Explorer and check the assembly list I get: We can see here that .NET has not actually loaded any of the dependencies of the Westwind.Utilities assembly. Also not loaded is the top level System.Data reference even though it's in the dependent assembly list of the top level project. Since this particular function I called only uses core System functionality (contained in mscorlib) there's in fact nothing else loaded beyond the main application and my Westwind.Utilities assembly that contains the method accessed. None of the dependencies of Westwind.Utilities loaded. If you were to open the assembly in a disassembler like Reflector or ILSpy, you would however see all the compiled in dependencies. The referenced assemblies are in the dependency list and they are loadable, but they are not immediately loaded by the application. In other words the C# compiler and .NET linker are smart enough to figure out the dependencies based on the code that actually is referenced from your application and any dependencies cascading down into the dependencies from your top level application into the referenced assemblies. In the example above the usage requirement is pretty obvious since I'm only calling a single static method and then exiting the app, but in more complex applications these dependency relationships become very complicated - however it's all taken care of by the compiler and linker figuring out what types and members are actually referenced and including only those assemblies that are in fact referenced in your code or required by any of your dependencies. The good news here is: That if you are referencing an assembly that has a dependency on something like System.Web in a few places that are not actually accessed by any of your code or any dependent assembly code that you are calling, that assembly is never loaded into memory! Some Hosting Environments pre-load Assemblies The load behavior can vary however. In Console and desktop applications we have full control over assembly loading so we see the core CLR behavior. However other environments like ASP.NET for example will preload referenced assemblies explicitly as part of the startup process - primarily to minimize load conflicts. Specifically ASP.NET pre-loads all assemblies referenced in the assembly list and the /bin folder. So in Web applications it definitely pays to minimize your top level assemblies if they are not used. Understanding when Assemblies Load To clarify and see it actually happen what I described in the first example , let's look at a couple of other scenarios. To see assemblies loading at runtime in real time lets create a utility function to print out loaded assemblies to the console: public static void PrintAssemblies() { var assemblies = AppDomain.CurrentDomain.GetAssemblies(); foreach (var assembly in assemblies) { Console.WriteLine(assembly.GetName()); } } Now let's look at the first scenario where I have class method that references internally uses System.Web. In the first scenario lets add a method to my main program like this: static void Main(string[] args) { Console.WriteLine(StringUtils.NewStringId()); Console.ReadLine(); PrintAssemblies(); } public static void WebLogEntry() { var entry = new WebLogEntry(); entry.UpdateFromRequest(); Console.WriteLine(entry.QueryString); } UpdateFromWebRequest() internally accesses HttpContext.Current to read some information of the ASP.NET Request object so it clearly needs a reference System.Web to work. In this first example, the method that holds the calling code is never called, but exists as a static method that can potentially be called externally at some point. What do you think will happen here with the assembly loading? Will System.Web load in this example? No - it doesn't. Because the WebLogEntry() method is never called by the mainline application (or anywhere else) System.Web is not loaded. .NET dynamically loads assemblies as code that needs it is called. No code references the WebLogEntry() method and so System.Web is never loaded. Next, let's add the call to this method, which should trigger System.Web to be loaded because a dependency exists. Let's change the code to: static void Main(string[] args) { Console.WriteLine(StringUtils.NewStringId()); Console.WriteLine("--- Before:"); PrintAssemblies(); WebLogEntry(); Console.WriteLine("--- After:"); PrintAssemblies(); Console.ReadLine(); } public static void WebLogEntry() { var entry = new WebLogEntry(); entry.UpdateFromRequest(); Console.WriteLine(entry.QueryString); } Looking at the code now, when do you think System.Web will be loaded? Will the before list include it? Yup System.Web gets loaded, but only after it's actually referenced. In fact, just until before the call to UpdateFromRequest() System.Web is not loaded - it only loads when the method is actually called and requires the reference in the executing code. Moral of the Story So what have we learned - or maybe remembered again? Dependent Assembly References are not pre-loaded when an application starts (by default) Dependent Assemblies that are not referenced by executing code are never loaded Dependent Assemblies are just in time loaded when first referenced in code All of this is nothing new - .NET has always worked like this. But it's good to have a refresher now and then and go through the exercise of seeing it work in action. It's not one of those things we think about everyday, and as I found out last week, I couldn't remember exactly how it worked since it's been so long since I've learned about this. And apparently I'm not the only one as several other people I had discussions with in relation to loaded assemblies also didn't recall exactly what should happen or assumed incorrectly that just having a reference automatically loads the assembly. The moral of the story for me is: Trying at all costs to eliminate an assembly reference from a component is not quite as important as it's often made out to be. For example, the Westwind.Utilities module described above has a logging component, including a Web specific logging entry that supports pulling information from the active HTTP Context. Adding that feature requires a reference to System.Web. Should I worry about this in the scope of this library? Probably not, because if I don't use that one class of nearly a hundred, System.Web never gets pulled into the parent process. IOW, System.Web only loads when I use that specific feature and if I am, well I clearly have to be running in a Web environment anyway to use it realistically. The alternative would be considerably uglier: Pulling out the WebLogEntry class and sticking it into another assembly and breaking up the logging code. In this case - definitely not worth it. So, .NET definitely goes through some pretty nifty optimizations to ensure that it loads only what it needs and in most cases you can just rely on .NET to do the right thing. Sometimes though assembly loading can go wrong (especially when signed and versioned local assemblies are involved), but that's subject for a whole other post…© Rick Strahl, West Wind Technologies, 2005-2012Posted in .NET  CSharp   Tweet !function(d,s,id){var js,fjs=d.getElementsByTagName(s)[0];if(!d.getElementById(id)){js=d.createElement(s);js.id=id;js.src="//platform.twitter.com/widgets.js";fjs.parentNode.insertBefore(js,fjs);}}(document,"script","twitter-wjs"); (function() { var po = document.createElement('script'); po.type = 'text/javascript'; po.async = true; po.src = 'https://apis.google.com/js/plusone.js'; var s = document.getElementsByTagName('script')[0]; s.parentNode.insertBefore(po, s); })();

  Read the article

• Box 2d basic questions

  - by philipp

  I am a bit new to box2d and I am developing an game with type and letters. I am using an svg font and generate the box2d bodies direct from the glyphs path definition, using the convex hull of them. I also have an decomposition routine the decomposes this hull if necessary. All this it is more or less working, except that I got some strange errors which definitely are caused by the scale factors. The problem is caused by two factors: first: the world scale of box2d, second: the the precision of curve-approximation of the glyph vectors. So through scaling down the input vertices for box2d, it happens that they become equal caused by numerical precision, what causes errors in box2d. Through scaling the my glyphs a bit up, this goes away. I also goes away if I chose a different world scale factor, but this slows down the whole animation quite much! So if my view port is about 990px * 600px and i want to animate Glyphs in box2d which should have a size from about 50px * 50px up to 300px * 300px, which scale factor of the b2world should i choose? How small should the smallest distance from on vertex to another be, while approximating the glyph vectors? Thanks for help greetings philipp EDIT:: I continued reading the docs of box2d and after rethinking of the units system, which is designed to handle object from 0.1 up to 10 meters, I calculated a scale factor of 75. So Objects 600px width will are 8 meters wide in box2d and even small objects of about 20px width will become 0.26 meters width in box2d. I will go on trying with this values, but if there is somebody out there who could give me a clever advice i would be happy!

  Read the article

• 2 way SSL between SOA and OSB

  - by Johnny Shum

  If you have a need to use 2 way SSL between SOA composite and external partner links, you can follow these steps. Create the identity keystores, trust keystores, and server certificates. Setup keystores and SSL on WebLogic Setup server to use 2 way SSL Configure your SOA composite's partner link to use 2 way SSL Configure SOA engine two ways SSL In this case,  I use SOA and OSB for the test.  I started with a separate OSB and SOA domains.  I deployed two soap based proxies on OSB and two composites on SOA.  In SOA, one composite invokes a OSB proxy service, the other is invoked by the OSB.  Similarly,  in OSB,  one proxy invokes a SOA composite and the other is invoked by SOA. 1. Create the identity keystores, trust keystores and the server certificates Since this is a development environment, I use JDK's keytool to create the stores and use self signing certificate.  For production environment, you should use certificates from a trusted certificate authority like Verisign.    I created a script below to show what is needed in this step.  The only requirement is when creating the SOA identity certificate, you MUST use the alias mykey. STOREPASS=welcome1KEYPASS=welcome1# generate identity keystore for soa and osb.  Note: For SOA, you MUST use alias mykeyecho "creating stores"keytool -genkey -alias mykey -keyalg "RSA" -sigalg "SHA1withRSA" -dname "CN=soa, C=US" -keystore soa-default-keystore.jks -storepass $STOREPASS -keypass $KEYPASS keytool -genkey -alias osbkey -keyalg "RSA" -sigalg "SHA1withRSA" -dname "CN=osb, C=US" -keystore osb-default-keystore.jks -storepass $STOREPASS -keypass $KEYPASS# listing keystore contentsecho "listing stores contents"keytool -list -alias mykey -keystore soa-default-keystore.jks -storepass $STOREPASSkeytool -list -alias osbkey -keystore osb-default-keystore.jks -storepass $STOREPASS# exporting certs from storesecho "export certs from  stores"keytool -exportcert -alias mykey -keystore soa-default-keystore.jks -storepass $STOREPASS -file soacert.derkeytool -exportcert -alias osbkey -keystore osb-default-keystore.jks -storepass $STOREPASS -file osbcert.der # import certs to trust storesecho "import certs"keytool -importcert -alias osbkey -keystore soa-trust-keystore.jks -storepass $STOREPASS -file osbcert.der -keypass $KEYPASSkeytool -importcert -alias mykey -keystore osb-trust-keystore.jks -storepass $STOREPASS -file soacert.der  -keypass $KEYPASS SOA suite uses the JDK's SSL implementation for outbound traffic instead of the WebLogic's implementation.  You will need to import the partner's public cert into the trusted keystore used by SOA.  The default trusted keystore for SOA is DemoTrust.jks and it is located in $MW_HOME/wlserver_10.3/server/lib.   (This is set in the startup script -Djavax.net.ssl.trustStore).   If you use your own trusted keystore, then you will need to import it into your own trusted keystore. keytool -importcert -alias osbkey -keystore $MW_HOME/wlserver_10.3/server/lib/DemoTrust.jks -storepass DemoTrustKeyStorePassPhrase  -file osbcert.der -keypass $KEYPASS If you do not perform this step, you will encounter this exception in runtime when SOA invokes OSB service using 2 way SSL Message send failed: sun.security.validator.ValidatorException: PKIX path building failed: sun.security.provider.certpath.SunCertPathBuilderException: unable to find valid certification path to requested target  2.  Setup keystores and SSL on WebLogic First, you will need to login to the WebLogic console, navigate to the server's configuration->Keystore's tab.   Change the Keystores type to Custom Identity and Custom Trust and enter the rest of the fields. Then you navigate to the SSL tab, enter the fields in the identity section and expand the Advanced section.  Since I am using self signing cert on my VM enviornment, I disabled Hostname verification.  In real production system, this should not be the case.   I also enabled the option "Use Server Certs", so that the application uses the server cert to initiate https traffic (it is important to enable this in OSB). Last, you enable SSL listening port in the Server's configuration->General tab. 3.  Setup server to use 2 way SSL If you follow the screen shot in previous step, you can see in the Server->Configuration->SSL->Advanced section, there is an option for Two Way Client Cert Behavior,  you should set this to Client Certs Requested and Enforced. Repeat step 2 and 3 done on OSB.  After all these configurations,  you have to restart all the servers. 4.  Configure your SOA composite's partner link to use 2 way SSL You do this by modifying the composite.xml in your project, locate the partner's link reference and add the property oracle.soa.two.way.ssl.enabled.   <reference name="callosb" ui:wsdlLocation="helloword.wsdl">    <interface.wsdl interface="http://www.examples.com/wsdl/HelloService.wsdl#wsdl.interface(Hello_PortType)"/>    <binding.ws port="http://www.examples.com/wsdl/HelloService.wsdl#wsdl.endpoint(Hello_Service/Hello_Port)"                location="helloword.wsdl" soapVersion="1.1">      <property name="weblogic.wsee.wsat.transaction.flowOption"                type="xs:string" many="false">WSDLDriven</property>   <property name="oracle.soa.two.way.ssl.enabled">true</property>    </binding.ws>  </reference> In OSB, you should have checked the HTTPS required flag in the proxy's transport configuration.  After this,  rebuilt the composite jar file and ready to deploy in the EM console later. 5.  Configure SOA engine two ways SSL Oracle SOA Suite uses both Oracle WebLogic Server and Sun Secure Socket Layer (SSL) stacks for two-way SSL configurations. For the inbound web service bindings, Oracle SOA Suite uses the Oracle WebLogic Server infrastructure and, therefore, the Oracle WebLogic Server libraries for SSL.  This is already done by step 2 and 3 in the previous section. For the outbound web service bindings, Oracle SOA Suite uses JRF HttpClient and, therefore, the Sun JDK libraries for SSL.  You do this by configuring the SOA Engine in the Enterprise Manager Console, select soa-infra->SOA Administration->Common Properties Then click at the link at the bottom of the page:  "More SOA Infra Advances Infrastructure Configuration Properties" and then enter the full path of soa identity keystore in the value field of the KeyStoreLocation attribute.  Click Apply and Return then navigate to the domain->security->credential. Here, you provide the password to the keystore.  Note: the alias of the certficate must be mykey as described in step 1, so you only need to provide the password to the identity keystore.   You accomplish this by: Click Create Map In the Map Name field, enter SOA, and click OK Click Create Key Enter the following details where the password is the password for the SOA identity keystore. 6.  Test and Trouble Shooting Once the setup is complete and server restarted, you can deploy the composite in the EM console and test it.  In case of error,  you can read the server log file to determine the cause of the error.  For example, If you have not setup step 5 and test 2 way SSL, you will see this in the log when invoking OSB from BPEL: java.lang.Exception: oracle.sysman.emSDK.webservices.wsdlapi.SoapTestException: oracle.fabric.common.FabricInvocationException: Unable to access the following endpoint(s): https://localhost.localdomain:7002/default/helloword ####<Sep 22, 2012 2:07:37 PM CDT> <Error> <oracle.soa.bpel.engine.ws> <rhel55> <AdminServer> <[ACTIVE] ExecuteThread: '1' for queue: 'weblogic.kernel.Default (self-tuning)'> <<anonymous>> <BEA1-0AFDAEF20610F8FD89C5> ............ <11d1def534ea1be0:-4034173:139ef56d9f0:-8000-00000000000002ec> <1348340857956> <BEA-000000> <got FabricInvocationException sun.security.provider.certpath.SunCertPathBuilderException: unable to find valid certification path to requested target If you have not enable WebLogic SSL to use server certificate in the console and invoke SOA composite from OSB using two ways SSL, you will see this error: ####<Sep 22, 2012 2:07:37 PM CDT> <Warning> <Security> <rhel55> <AdminServer> <[ACTIVE] ExecuteThread: '6' for queue: 'weblogic.kernel.Default (self-tuning)'> <<WLS Kernel>> <> <11d1def534ea1be0:-51f5c76a:139ef5e1e1a:-8000-00000000000000e2> <1348340857776> <BEA-090485> <CERTIFICATE_UNKNOWN alert was received from localhost.localdomain - 127.0.0.1. The peer has an unspecified issue with the certificate. SSL debug tracing should be enabled on the peer to determine what the issue is.> ####<Sep 22, 2012 2:07:37 PM CDT> <Warning> <Security> <rhel55> <AdminServer> <[ACTIVE] ExecuteThread: '6' for queue: 'weblogic.kernel.Default (self-tuning)'> <<WLS Kernel>> <> <11d1def534ea1be0:-51f5c76a:139ef5e1e1a:-8000-00000000000000e4> <1348340857786> <BEA-090485> <CERTIFICATE_UNKNOWN alert was received from localhost.localdomain - 127.0.0.1. The peer has an unspecified issue with the certificate. SSL debug tracing should be enabled on the peer to determine what the issue is.> ####<Sep 22, 2012 2:27:21 PM CDT> <Warning> <Security> <rhel55> <AdminServer> <[ACTIVE] ExecuteThread: '0' for queue: 'weblogic.kernel.Default (self-tuning)'> <<anonymous>> <> <11d1def534ea1be0:-51f5c76a:139ef5e1e1a:-8000-0000000000000124> <1348342041926> <BEA-090497> <HANDSHAKE_FAILURE alert received from localhost - 127.0.0.1. Check both sides of the SSL configuration for mismatches in supported ciphers, supported protocol versions, trusted CAs, and hostname verification settings.> References http://docs.oracle.com/cd/E23943_01/admin.1111/e10226/soacompapp_secure.htm#CHDCFABB   Section 5.6.4 http://docs.oracle.com/cd/E23943_01/web.1111/e13707/ssl.htm#i1200848

  Read the article

• Uploading documents to WSS (Windows Sharepoint Services) using SSIS

  - by Randy Aldrich Paulo

  Recently I was tasked to create an SSIS application that will query a database, split the results with certain criteria and create CSV file for every result and upload the file to a Sharepoint Document Library site. I've search the web and compiled the steps I've taken to build the solution. Summary: A) Create a proxy class of WSS Copy.asmx. B) Create a wrapper class for the proxy class and add a mechanism to check if the file is existing and delete method. C) Create an SSIS and call the wrapper class to transfer the files.   A) Creating Proxy Class 1) Go to Visual Studio Command Prompt type wsdl http://[sharepoint site]/_vti_bin/Copy.asmx this will generate the proxy class (Copy.cs) that will be added to the solution. 2) Add Copy.cs to solution and create another constructor for Copy() that will accept additional parameters url, userName, password and domain.   public Copy(string url, string userName, string password, string domain) { this.Url = url; this.Credentials = new System.Net.NetworkCredential(userName, password, domain); } 3) Add a namespace.     B) Wrapper Class Create a C# new library that references the Proxy Class.         C) Create SSIS SSIS solution is composed of:   1) Execute SQL Task, returns a single column rows containing the criteria. 2) Foreach Loop Container - loops per result from query (SQL Task) and creates a CSV file on a certain folder. 3) Script Task - calls the wrapper class to upload CSV files located on a certain folder to targer WSS Document Library Note: I've created another overload of CopyFiles that accepts a Directory Info instead of file location that loops thru the contents of the folder. Designer View Variable View

  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

• SQL SERVER – SELECT TOP Shortcut in SQL Server Management Studio (SSMS)

  - by pinaldave

  This is tool is pretty old, yet always comes as a handy tip. I had a great trip at TechEd in India. And, during one of my presentations, I was asked if there are any shortcuts to SELECT only TOP 100 records from SSMS. I immediately told him that if he explores the table in SSMS, he can just right click on it and SELECT TOP 1000 records. If he wanted only 100 records, then he could edit that 1000 to 100 by means of going to Options. Go to Options, then hover the mouse over the SQL Server Object Explorer, then proceed to Commands. Afterwards, change the Value for Select Top <n> Audit Records. After narrating the steps, he told me that he was not looking for the right click option; rather he was asking if there is any kind of keyboard shortcut for convenience’s sake. Actually, a keyboard shortcut is also possible. SQL Server Management Studio (SSMS) lets you configure the settings you want using a shortcut. Here is how you can do it. Go to Options, then to Environment. Proceed to Keyboard, and from there, configure your T-SQL with the desired keyword. Now, open SSMS New Query Window, and then click and type in any table name.  After that, just hit the shortcut you just made earlier. Doing this should display TOP 100 records in the Result window. I am sure this trick is quite old, but it is still helpful to many. Reference: Pinal Dave (http://blog.SQLAuthority.com) Filed under: Pinal Dave, SQL, SQL Add-On, SQL Authority, SQL Query, SQL Server, SQL Tips and Tricks, T SQL, Technology

  Read the article

• 2D grid with multiple types of objects

  - by Alexandre P. Levasseur

  This is my first post here in programmers.stackexchange (I'm a regular on SO). I hope this isn't too general. I'm trying a simple project to learn Java from something I've seen done in the past. Basically, it's an AI simulation where there are herbivorous and carnivorous creatures and both must try to survive. The part I am trying to come up with is that of the board itself. Let's assume very simple rules. The board must be of size X by Y and only one element can be in one place at one time. For example, a critter cannot be in the same tile as a food block. There can be obstacles (rocks, trees..), there can be food, there can be critters of any type. Assuming these rules, what would be one good way to represent this situation ? This is what I came up with and want suggestions if possible: Use multiple levels of inheritance to represent all the different possible objects (AbstractObject - (NonMovingObject - (Food, Obstacle) , MovingObject - Critter - (Carnivorous, Herbivorous))) and use polymorphism in a 2D array to store the instances and still have access to lower level methods. Many thanks. Edit: Here is the graphic representation of the structure I have in mind.

  Read the article

• Customize the Windows Media Center Start Menu with Media Center Studio

  - by DigitalGeekery

  Do you ever wish you could change the WMC start menu? Maybe move some of the tiles and strips around to different locations, add new ones, or eliminate some altogether? Today we look at how to do it using Media Center Studio. Download and install Media Center Studio. (Download link below) You’ll also want to make sure you have Windows Media Center closed before running Media Center Studio. Many of the actions cannot be performed with Media Center open. Once installed, you can open Media Center Studio from the Windows Start Menu. When you first open Media Center Studio you’ll be on the Themes tab. Click on the Start Menu tab. It should be noted that Media Center Studio is a Beta application, and it did crash on us a few times, so it’s a good idea to save your work frequently. You can save your changes by selecting Save on the Home tab, or by clicking the small disk icon at the top left. We also found that that trying to launch Media Center from the Start Media Center button on the application ribbon typically didn’t work. Opening Windows Media Center from the Windows Start Menu is preferred.   When you’re on the Start Menu tab you will see the Windows Media Center menu strips and tiles. Click the arrows located at the right, left, top, and bottom of the screen to scroll through the various menu strips.   Hiding and Removing Tiles and Menu Strips. If there is an entire menu strip that you never use and would like to remove from Media Center, simply uncheck the box to the left of the the title above that menu strip. If you’d like to hide individual tiles, uncheck the box next to the name of the individual tile. Renaming Tiles and Strips To rename a tile or menu strip, click on the small notepad icon next to the title. Note: If you do not see a small notepad icon next to the title, then the title is not editable. This applies to many of the “Promo” tiles. The title will turn into a text input box so that you can edit the name. Click away from the text box when finished. Here we will change the title of the default Movie strip to “Flicks.” Change the Default Tile and Menu Strip The Default menu strip is the strip that is highlighted, or on focus, when you open Media Center.   To change the default strip, simply click once on another strip to highlight it, and then save your work. In our example, I’m going to make our newly renamed “Flicks” strip the default.   Each menu strip has a default tile. This is the tile that is active, or on focus, when you select the menu strip. To change the default tile on a strip, click once on the tile. You will see it outlined in light blue. Now just simply save your changes. In our example below, we’ve changed the default tile on the TV strip to “guide.”   Moving Tiles and Menu Strips You can move an entire Menu Strip up or down on the screen. When you hover your mouse over the a menu strip, you will see up and down arrows appear to the right and left of the title. Click on the arrows to move the strip up or down.   You will see the menu strip appear in it’s new position.   To move a tile to a new menu strip, click and drag the tile you’d like to move. When you begin to drag the tile, green plus (+) signs will appear in between the tiles. Drag and drop the tile onto to any of these green plus signs to move it to that location. When you’ve dragged the tile over an acceptable position, you’ll see the  red “Move” label next to your cursor turn to a blue “Move to” label. Now you can drop the tile into position. You’ll see the tile located in it’s new position.   Adding a New Custom Menu Strip Click on the Start Menu tab and then select the Menu Strip button.   You will see a new Custom Menu strip appear on your Start Menu with the default name of Custom menu. You can change the name by clicking on the notepad icon just as we did earlier. For our example, we’ll change the name of the new strip to Add-ins. To add a new tile, click on Entry Points at the lower left of the application window. This will reveal all of your available Entry Points that can be added to the Media Center Menu. You should see the built-in Media Center Games and any Media Center Plug-ins you have added to your system. You can then drag and drop any of the Entry Points onto any of the Menu Strips. Below we’ve added Media Browser to our custom Add-ins menu strip. You can also add additional applications to launch directly from Media Center. Click on the Application button on the Start Menu tab. Note: Many applications may not work with your remote, but with keyboard and mouse only.    Type in a title which will appear under the tile in Media Center, and then type the path to the application. In our example, we will add Internet Explorer 8. Note: Be sure to add the actual path to the application and not just a link on the desktop. Click any of the check boxes to select any options under Required Capabilities. You can also browse to choose an image if you don’t care for the image that appears automatically.   Next, you can select keyboard strokes to press to exit the application and return to Media Center. Click the green plus (+) button. When prompted, press a key you’ll use to close the program. Repeat the process if you’d also like to select a keystroke to kill the program.   You’ll see your button programs listed below. When you’re finished, save your work and close out of Media Center Studio.   Now your new program entry point will appear in the Entry Points section. Drag the icon to the desired position on the Start Menu and save again before exiting Media Center Studio. When you open Media Center you will see your new application on the start menu. Click the tile to open the application just as you would any other tile. The application will open and minimize Media Center. When you press the key you choose to close the program, Windows Media Center will automatically be restored. Note: You can also exit the application through normal methods by clicking the red “X” or File > Exit. Conclusion Media Center Studio is a Beta application which the developer freely admits still has some bugs. Despite it’s flaws Media Center Studio is a powerful tool, and when it comes to customizing your Media Center start menu, it’s pretty much the only game in town. It works with both Vista and Windows 7, and according to the developer, has not been officially tested with extenders. Media Center Studio can also be used to add custom themes to Windows 7 Media Center and we’ll be covering that in a future article. Looking for more ways to customize your Media Center experience? Be sure to check out our earlier posts on Media Browser, as well as how to add Hulu, Boxee, and weather conditions your Windows 7 Media Center. Download Media Center Studio Similar Articles Productive Geek Tips Using Netflix Watchnow in Windows Vista Media Center (Gmedia)How To Rip a Music CD in Windows 7 Media CenterSchedule Updates for Windows Media CenterStartup Customizations for Media Center in Windows 7Automatically Start Windows 7 Media Center in Live TV Mode TouchFreeze Alternative in AutoHotkey The Icy Undertow Desktop Windows Home Server – Backup to LAN The Clear & Clean Desktop Use This Bookmarklet to Easily Get Albums Use AutoHotkey to Assign a Hotkey to a Specific Window Latest Software Reviews Tinyhacker Random Tips Acronis Online Backup DVDFab 6 Revo Uninstaller Pro Registry Mechanic 9 for Windows Video Toolbox is a Superb Online Video Editor Fun with 47 charts and graphs Tomorrow is Mother’s Day Check the Average Speed of YouTube Videos You’ve Watched OutlookStatView Scans and Displays General Usage Statistics How to Add Exceptions to the Windows Firewall

  Read the article

• I can't finish installation of 12.10 on Windows 8

  - by Janna Zhou

  When I followed the instruction to install Ubuntu 12.10 with wubi, the come out window ask me to reboot the computer to finish the installation. I followed the instruction and stopped with the message I have left below. Could you please tell me how to finish installation? I'm worried if I follow the instruction below to remove_hiberfile in Windows it would result in a failure to boot my Window 8 system: Completing the Ubuntu installation. For more installation boot options, press 'ESC'now... 0 Busy Box v1.19.3 (Ubuntu 1:1.19.3-7ubuntu1) built-in shell (ash) Enter 'help' for a list of built-in commands. (initramfs) windows is hibernated, refused to mount. Failed to mount '/dev/sda2/: Operation not permitted the NTFS partition is hibernated. Please resume and shutdown Windows properly, or mount the volume read-only with the 'ro' mount option, or mount the volume read-write with the 'remove_hiberfile' mount option. For example type on the command line: mount -t ntsfs-3g -0 remove_hiberfile/dev/sda2/isodevice mount:mounting/dev/sda2 on /isodevice failed: No such device Could not find the ISO/ubuntu/install/installation.iso This could also happen if the file system is not clean because of an operating system crash, an interrupted boot process, an improper shutdown, or unplugging of a removable device without first unmounting or ejecting it. To fix this, simply reboot into windows, let it fully start, log in, run 'chkdsk/r', then gracefully shut down and reboot back into Windows. After this you should be able to reboot again and resume the installation.

  Read the article

< Previous Page | 972 973 974 975 976 977 978 979 980 981 982 983  | Next Page >