Search Results

Search found 9814 results on 393 pages for 'personal technology'.

Page 367/393 | < Previous Page | 363 364 365 366 367 368 369 370 371 372 373 374  | Next Page >

• Second HDD not seen by Windows 7 on Dell Xps l501x

  - by George

  I have a Dell XPS Laptop (l501x). I have replaced the original Seagate 500GB hard drive with an SSD Intel 320 120GB when I first purchased it a year ago. It's been working great. The laptop is booting in about 23 seconds, so the SSD is great. I have an Acronis image created that I came back to every three months just to keep everything clean. The SSD is partitioned with one logical drive for my data. So, recently I thought since I am not using my optical drive often enough to swap it out with a HDD caddy and add my seagate 500gb hard drive. I ordered the caddy placed the HDD in it and now load Windows. It just hangs on the screen that should show the Windows logo. I have tried everything that I know and searched it online. I have uninstalled the SATA controller AHCI and let Windows install it. Still will not boot into windows. I wanted to mention that the Seagate 500GB drive was the one that came with my laptop before I switched to the Intel SSD. As you know Intel has this application called Intel Rapid Technology which loads once in a while and shows the second hard drive, but then, when I restart it hangs again and Windows will not load. As soon as I remove the HDD Caddy and restart it loads Windows fine. I also formated the Seagate 500GB HDD in NTFS and still Windows will not load. When I go into the BIOS it shows the Fixed SSD and also "Sata ODD 500GB" instead of the optical drive but it will not boot into Windows when the HDD caddy is present. There is nothing wrong with the caddy. I have another laptop (Asus) and I installed the HDD caddy and Windows 7 loads without any glitch. I don't get it. I have also flashed the BIOS because Dell had a new version (A08). I also wanted to add that I refreshed Disk Management and the Device Manager and the second drive does not display. At this point I think it's a Windows issue so before I reinstall Windows 7 Home Premium from scratch I wanted to see if there was anything I was missing. Any advice would be greatly appreciated.

  Read the article

• What Sort of Server Setup Am I Likely to Need? - School A/V streaming

  - by DeathMagus

  My prior experience with servers has generally been limited to home file-sharing servers, low-traffic web-servers, and the like. This leaves me with the technical knowledge of how to set up a system, but little experience in terms of scaling said system. My current project, however, has me as the technical lead in setting up a school for online audio and video streaming. The difficulty I'm running into is that I don't quite have the experience to guess what they'll need, and they don't have the experience to tell me - so I've tried to ask as many pertinent questions about what they want to do with their server, and here's what I found out: About 1000 simultaneous users, and hoping to expand (possibly significantly) Both video and audio streaming, at obviously the highest quality possible Support for both live and playlist-based streaming. Probably only one channel, but as it's an educational opportunity, I imagine letting them have a few more wouldn't hurt. No word on whether they're locked into Windows or whether Linux is acceptable. Approximate budget - $7000. It may actually be about $2k less than this, because of a mishap with another technology firm (they ordered a $7000 DV tape deck for some reason, and now the company wants them to pay a 30% restocking fee). The tentative decisions I've already made: I'm planning on using Icecast 2 for my streaming server, fed by VLC Shoutcast encoding. Since the school already has a DMZ set up, I plan on placing the Icecast server in there, and feeding it through their intranet from a simple workstation computer in their studios. This system isn't in any way mission critical - it's an education tool (they're a media magnet school), so I figure redundancy is not worthwhile to them from a cost:benefit perspective. What I don't know is this: How powerful of a server will I need? What is likely to be my major throttle - bandwidth? How can I mitigate that? Will I need anything special for the encoding workstation other than professional video and audio capture cards and a copy of VLC? Are there any other considerations that I'm simply missing? Thanks a lot for any help - if there's more information you need, let me know and I'll tell you all I can.

  Read the article

• Can Remote Desktop Services be deployed and administered by PowerShell alone, without a Domain in WIndows Server 2012 and 2012 R2?

  - by Warren P

  Windows Server 2008 R2 allowed deployment of Terminal Server (Remote Desktop Services) without a domain, and without any insistence on domains. This was very useful, especially for standalone virtual or cloud deployments of a server that is managed remotely for a remote client who has no need or desire for any ActiveDirectory or Domain features. This has become steadily more and more difficult as Microsoft restricts its technologies further and further in each Windows release. With Windows Server 2012, configuring licensing for Remote Desktop Services, is more difficult when not on a domain, but possible still. With Windows Server 2012 R2 (at least in the preview) the barriers are now severe: The Add/Remove Roles and Features wizard in Windows Server 2012 R2 has a special RDS deployment mode that has a rule that says if you aren't on a domain you can't deploy. It tells you to create or join a domain first. This of course comes in direct conflict with the fact that an Active Directory domain controller should not be the same machine as a terminal server machine. So Microsoft's technology is not such much a Cloud Operating System as a Cluster of Unwanted Nodes, needed to support the one machine I actually WANT to deploy. This is gross, and so I am trying to find a workaround. However if you skip that wizard and just go check the checkboxes in the main Roles/Features wizard, you can deploy the features, but the UI is not there to configure them, and when you go back to the RDS configuration page on the roles wizard, you get a message saying you can not administer your Remote Desktop Services system when you are logged in as a Local-Computer Administrator, because although you have all admin priveleges you could have (in your workgroup based system), the RDS configuration UI will not accept those credentials and let you continue. My question in brief is, can I still somehow, obtain the following end result: I need to allow 10-20 users per system to have an RDS (TS) session. I do not need any of the fancy pants RDS options, unless Microsoft somehow depends on those features being present. I believe I need the "RDS Session Host" as this is the guts of "Terminal Server". Microsoft says it is "full Windows desktop for Remote Desktop Services client. I need to configure licensing so that the Grace Period does not expire leaving my RDS non functional, so this probably means I need a way to configure TS CALs. If all of the above could technically be done with the judicious use of the PowerShell, I am prepared to even consider developing all the PowerShell scripts I would need to do the above. I'm not asking someone to write that for me. What I'm asking is, does anyone know if there is a technical impediment to what I want to do above, other than the deliberate crippling of the 2012 R2 UI for Workgroup users? Would the underlying technologies all still work if I manipulate and control them from a PowerShell script? Obviously a 1 word Yes or No answer isn't that useful to anyone, so the question is really, yes or no, and why? In the case the answer is Yes, then how.

  Read the article

• MS DPM 2007: Testing the Recovery for a Production Domain

  - by NewToDPM

  Hi everybody! MS DPM 2007 is a new technology in my company, and so am I to the product. We have a classic Microsoft domain with two DCs, Exchange 2007 and a couple Web/MS SQL servers. I have deployed DPM one month ago on the domain, and after fixing the various issues I got with the replicas inconsistence and adapting the schedule and retention range to the server storage pool size, I can say the backup system is working correctly (no errors) as of today. However, there is one problem: we did not attempt to restore from the backups yet, which is a big no-no of course. I'm not sure about the way I should handle this, my main concern being Exchange and the System State of the DCs. From my understanding, DPM can only protect AND restore data on a server which is part of the same domain as the backup server. If I restore the System State (containing Active Directory) and the Exchange Storage Groups on a testing server, I am afraid it would completely disturb the domain functioning (for example, having two primary DCs on the domain). I am thinking about building a second DPM server on a testing separate domain which would mirror the replicas and then restore it on testing servers from this new domain. Is it the right way to handle the data recovery testing? How did you do on your domain when you first deployed DPM? I'd be grateful for any link/documentation or advice. Thank you in advance for your help! EDIT: Two options seem possible so far: i. Create another DC/Exchange server in the alternate location; ii. Create a separate domain in the alternate location and setup a trust between this domain and the production one. The option i is certainly the best but implies setting up a secondary Exchange server, with a dedicated public IP address so that if Exchange #1 dies, we can still send emails with Exchange #2. I don't know how complex this can be and would need to discuss it with my colleagues. The option ii would only fit the testing purposes. My only question regarding this is: if my production and DPM servers are part of domain A, and there is a trust between domains A and B, can I restore a domain A content to any domain B server?

  Read the article

• Can not get sound over hdmi in kubuntu 9.10

  - by user32509

  I have used a hdmi cable to connect my lcd (which is connected with my speakers) with my nvida 275 gtx grafic card. I can not get the sound output to work. The hardware itself is working probably - I tested it under windows. Currently I am running Kubuntu 9.10 64 with Nvidia 190.53. The sound output worked fine before I installed the hdmi connection. (German output - i can change it, if you tell me how :)) aplay -l **** Liste von PLAYBACK Geräten **** Karte 0: Intel [HDA Intel], Gerät 0: ALC889A Analog [ALC889A Analog] Untergeordnete Geräte: 1/1 Untergeordnetes Gerät '0: subdevice #0 Karte 0: Intel [HDA Intel], Gerät 1: ALC889A Digital [ALC889A Digital] Untergeordnete Geräte: 1/1 Untergeordnetes Gerät '0: subdevice #0 aplay -L front:CARD=Intel,DEV=0 HDA Intel, ALC889A Analog Front speakers surround40:CARD=Intel,DEV=0 HDA Intel, ALC889A Analog 4.0 Surround output to Front and Rear speakers surround41:CARD=Intel,DEV=0 HDA Intel, ALC889A Analog 4.1 Surround output to Front, Rear and Subwoofer speakers surround50:CARD=Intel,DEV=0 HDA Intel, ALC889A Analog 5.0 Surround output to Front, Center and Rear speakers surround51:CARD=Intel,DEV=0 HDA Intel, ALC889A Analog 5.1 Surround output to Front, Center, Rear and Subwoofer speakers surround71:CARD=Intel,DEV=0 HDA Intel, ALC889A Analog 7.1 Surround output to Front, Center, Side, Rear and Woofer speakers iec958:CARD=Intel,DEV=0 HDA Intel, ALC889A Digital IEC958 (S/PDIF) Digital Audio Output null Discard all samples (playback) or generate zero samples (capture) pulse Playback/recording through the PulseAudio sound server And i disabled mute in kmix an all channels :) Edit: lspci -v ... 00:1b.0 Audio device: Intel Corporation 82801I (ICH9 Family) HD Audio Controller (rev 02) Subsystem: Giga-byte Technology Device a022 Flags: bus master, fast devsel, latency 0, IRQ 22 Memory at ea400000 (64-bit, non-prefetchable) [size=16K] Capabilities: [50] Power Management version 2 Capabilities: [60] Message Signalled Interrupts: Mask- 64bit+ Queue=0/0 Enable- Capabilities: [70] Express Root Complex Integrated Endpoint, MSI 00 Capabilities: [100] Virtual Channel <?> Capabilities: [130] Root Complex Link <?> Kernel driver in use: HDA Intel Kernel modules: snd-hda-intel ... cat /proc/asound/version Advanced Linux Sound Architecture Driver Version 1.0.20. lsmod | grep snd_hda_intel snd_hda_intel 31880 2 snd_hda_codec 87584 2 snd_hda_codec_realtek,snd_hda_intel snd_pcm 93160 3 snd_hda_intel,snd_hda_codec,snd_pcm_oss snd 77096 16 snd_hda_codec_realtek,snd_hda_intel,snd_hda_codec,snd_hwdep,snd_pcm_oss,snd_mixer_oss,snd_pcm,snd_seq_oss,snd_rawmidi,snd_seq,snd_timer,snd_seq_device snd_page_alloc 10928 2 snd_hda_intel,snd_pcm I think I am missing the something-hdmi module? Is there such a thing?

  Read the article

• Can I use the same machine as a client and server for SSH?

  - by achraf

  For development tests, I need to setup an SFTP server. So I want to know if it's possible to use the same machine as the client and the server. I tried and I keep getting this error: > Permission denied (publickey). > Connection closed and by running ssh -v agharroud@localhost i get : > OpenSSH_3.8.1p1,OpenSSL 0.9.7d 17 Mar > debug1: Reading configuration data /etc/ssh_config > debug1: Connecting to localhost [127.0.0.1] port 22. > debug1: Connection established. > debug1: identity file /home/agharroud/.ssh/identity type -1 > debug1: identity file /home/agharroud/.ssh/id_rsa type 1 > debug1: identity file /home/agharroud/.ssh/id_dsa type -1 > debug1: Remote protocol version 2.0, remote software version OpenSSH_3.8.1p1 > debug1: match: OpenSSH_3.8.1p1 pat OpenSSH* > debug1: Enabling compatibility mode for protocol 2.0 > debug1: Local version string SSH-2.0-OpenSSH_3.8.1p1 > debug1: SSH2_MSG_KEXINIT sent > debug1: SSH2_MSG_KEXINIT received > debug1: kex:server->client aes128-cbc hmac-md5 none > debug1: kex: client->server aes128-cbc hmac-md5 none > debug1: SSH2_MSG_KEX_DH_GEX_REQUEST(1024<1024<8192) sent > debug1: expecting SSH2_MSG_KEX_DH_GEX_GROUP > debug1: SSH2_MSG_KEX_DH_GEX_INIT sent > debug1: expecting SSH2_MSG_KEX_DH_GEX_REPLY > debug1: Host 'localhost' is known and matches the RSA host key. > debug1: Found key in /home/agharroud/.ssh/known_hosts:1 > debug1: ssh_rsa_verify: signature correct > debug1: SSH2_MSG_NEWKEYS sent > debug1: expecting SSH2_MSG_NEWKEYS > debug1: SSH2_MSG_NEWKEYS received > debug1: SSH2_MSG_SERVICE_REQUEST sent > debug1: SSH2_MSG_SERVICE_ACCEPT > received > > ****USAGE WARNING**** > > This is a private computer system. This computer system, including all > related equipment, networks, and network devices (specifically > including Internet access) are provided only for authorized use. This > computer system may be monitored for all lawful purposes, including to > ensure that its use is authorized, for management of the system, to > facilitate protection against unauthorized access, and to verify > security procedures, survivability, and operational security. Monitoring > includes active attacks by authorized entities to test or verify the > security of this system. During monitoring, information may be > examined, recorded, copied and used for authorized purposes. All > information, including personal information, placed or sent over this > system may be monitored. > > Use of this computer system, authorized or unauthorized, > constitutes consent to monitoring of this system. Unauthorized use may > subject you to criminal prosecution. Evidence of unauthorized use collected > during monitoring may be used for administrative, criminal, or other > adverse action. Use of this system constitutes consent to monitoring for > these purposes. > > debug1: Authentications that can continue: publickey > debug1: Next authentication method: publickey > debug1: Trying private key:/home/agharroud/.ssh/identity > debug1: Offering public key:/home/agharroud/.ssh/id_rsa > debug1:Authentications that can continue:publickey > debug1: Trying private key:/home/agharroud/.ssh/id_dsa > debug1: No more authentication methods to try. > Permission denied (publickey). Any ideas about the problem ? thanks !

  Read the article

• Most secure way to access my home Linux server while I am on the road? Specialized solution wanted

  - by Ace Paus

  I think many people may be in my situation. I travel on business with a laptop. And I need secure access to files from the office (which in my case is my home). The short version of my question: How can I make SSH/SFTP really secure when only one person needs to connect to the server from one laptop? In this situation, what special steps would make it almost impossible for anyone else to get online access to the server? A lot more details: I use Ubuntu Linux on both my laptop (KDE) and my home/office server. Connectivity is not a problem. I can tether to my phone's connection if needed. I need access to a large number of files (around 300 GB). I don't need all of them at once, but I don't know in advance which files I might need. These files contain confidential client info and personal info such as credit card numbers, so they must be secure. Given this, I don't want store all these files on Dropbox or Amazon AWS, or similar. I couldn't justify that cost anyway (Dropbox don't even publish prices for plans above 100 GB, and security is a concern). However, I am willing to spend some money on a proper solution. A VPN service, for example, might be part of the solution? Or other commercial services? I've heard about PogoPlug, but I don't know if there is a similar service that might address my security concerns? I could copy all my files to my laptop because it has the space. But then I have to sync between my home computer and my laptop and I found in the past that I'm not very good about doing this. And if my laptop is lost or stolen, my data would be on it. The laptop drive is an SSD and encryption solutions for SSD drives are not good. Therefore, it seems best to keep all my data on my Linux file server (which is safe at home). Is that a reasonable conclusion, or is anything connected to the Internet such a risk that I should just copy the data to the laptop (and maybe replace the SSD with an HDD, which reduces battery life and performance)? I view the risks of losing a laptop to be higher. I am not an obvious hacking target online. My home broadband is cable Internet, and it seems very reliable. So I want to know the best (reasonable) way to securely access my data (from my laptop) while on the road. I only need to access it from this one computer, although I may connect from either my phone's 3G/4G or via WiFi or some client's broadband, etc. So I won't know in advance which IP address I'll have. I am leaning toward a solution based on SSH and SFTP (or similar). SSH/SFTP would provided about all the functionality I anticipate needing. I would like to use SFTP and Dolphin to browse and download files. I'll use SSH and the terminal for anything else. My Linux file server is set up with OpenSSH. I think I have SSH relatively secured. I'm using Denyhosts too. But I want to go several steps further. I want to get the chances that anyone can get into my server as close to zero as possible while still allowing me to get access from the road. I'm not a sysadmin or programmer or real "superuser". I have to spend most of my time doing other things. I've heard about "port knocking" but I have never used it and I don't know how to implement it (although I'm willing to learn). I have already read a number of articles with titles such as: Top 20 OpenSSH Server Best Security Practices 20 Linux Server Hardening Security Tips Debian Linux Stop SSH User Hacking / Cracking Attacks with DenyHosts Software more... I have not implemented every single thing I've read about. I probably can't do that. But maybe there is something even better I can do in my situation because I only need access from a single laptop. I'm just one user. My server does not need to be accessible to the general public. Given all these facts, I'm hoping I can get some suggestions here that are within my capability to implement and that leverage these facts to create a great deal better security than general purpose suggestions in the articles above.

  Read the article

• How to install RAID drivers on already installed Windows 7?

  - by happysencha

  64-bit Windows 7 Ultimate 6GB RAM Intel i7 920 Intel X25-M SSD 80GB 2,5" Club 3D Radeon HD5750 GA-EX58-UD4P Motherboard I've been running fine with Windows 7 installed on the SSD. I wanted to create an mirrored Raid-1 setup for backups using two hard disks, so I ordered two Samsung HD203WI. This motherboard supports two different RAID controllers, the Intel's ICH10R and Gigabyte's SATA2 SATA controller. There are 6 SATA ports behind the ICH10R and 2 SATA ports for the Gigabyte controller. I googled around and seemed that the ICH10R is a better choice and since then I've been trying to make it work. When I activate the [RAID] mode from BIOS, the Windows 7 gives BSOD exactly as described by this guy: "Windows 7 will start to boot, it gets to the screen where there are 4 colors coming together and it blue screens and restarts no matter what I do." First thing I did: turned off the RAID and booted to Windows and tried to install the SATA RAID drivers from Gigabyte. I launch the driver installation program and it gives "This computer does not meet the minimum requirements for installing the software" error. I then tried Intel's Rapid Storage Technology drivers (which apparently is the same as the one offered at Gigabyte's site), but it resulted in exactly the same error. I then detached the new Samsung hard disks from the SATA ports, but left the [RAID] enabled in BIOS. To my surprise, it still BSOD'd, so at this point I knew it is an OS/driver issue. Also, I tried with the Gigabyte's RAID enabled (while the ICH10R RAID disabled) and it booted just fine. So then I thought, that maybe I can't install the RAID drivers from within the OS. So I caused the BSOD on purpose once again, and then with ICH10R RAID activated and Samsung hard disks attached, I choose the Windows 7 Recovery mode in the boot menu. It sees some problem(s), tries to repair, does not succeed and does not ask for drivers (which I put on a USB stick) to install. I also tried to use the command-line in the recovery: "rundll32 syssetup, SetupInfObjectInstallAction DefaultInstall 128 iaStor.inf" but it gave "Installation failed." So I'm clueless how should I proceed. Do I really need to re-install Windows 7 and load RAID drivers in the Win7 setup? I don't want to install any OS on the RAID, the Windows 7 is and will be on the SSD. I just want to have a RAID-1 backup using those two hard disks. I mean why would I need to re-install operating system to add RAID setup?

  Read the article

• boost fusion: strange problem depending on number of elements on a vector

  - by ChAoS

  I am trying to use Boost::Fusion (Boost v1.42.0) in a personal project. I get an interesting error with this code: #include "boost/fusion/include/sequence.hpp" #include "boost/fusion/include/make_vector.hpp" #include "boost/fusion/include/insert.hpp" #include "boost/fusion/include/invoke_procedure.hpp" #include "boost/fusion/include/make_vector.hpp" #include <iostream> class Class1 { public: typedef boost::fusion::vector<int,float,float,char,int,int> SequenceType; SequenceType s; Class1(SequenceType v):s(v){} }; class Class2 { public: Class2(){} void met(int a,float b ,float c ,char d ,int e,int f) { std::cout << a << " " << b << " " << c << " " << d << " " << e << std::endl; } }; int main(int argn, char**) { Class2 p; Class1 t(boost::fusion::make_vector(9,7.66f,8.99f,'s',7,6)); boost::fusion::begin(t.s); //OK boost::fusion::insert(t.s, boost::fusion::begin(t.s), &p); //OK boost::fusion::invoke_procedure(&Class2::met,boost::fusion::insert(t.s, boost::fusion::begin(t.s), &p)); //FAILS } It fails to compile (gcc 4.4.1): In file included from /home/thechaos/Escriptori/of_preRelease_v0061_linux_FAT/addons/ofxTableGestures/ext/boost/fusion/include/invoke_procedur e.hpp:10, from problema concepte.cpp:11: /home/thechaos/Escriptori/of_preRelease_v0061_linux_FAT/addons/ofxTableGestures/ext/boost/fusion/functional/invocation/invoke_procedure.hpp: I n function ‘void boost::fusion::invoke_procedure(Function, const Sequence&) [with Function = void (Class2::*)(int, float, float, char, int, in t), Sequence = boost::fusion::joint_view<boost::fusion::joint_view<boost::fusion::iterator_range<boost::fusion::vector_iterator<const boost::f usion::vector<int, float, float, char, int, int, boost::fusion::void_, boost::fusion::void_, boost::fusion::void_, boost::fusion::void_>, 0>, boost::fusion::vector_iterator<boost::fusion::vector<int, float, float, char, int, int, boost::fusion::void_, boost::fusion::void_, boost::fus ion::void_, boost::fusion::void_>, 0> >, const boost::fusion::single_view<Class2*> >, boost::fusion::iterator_range<boost::fusion::vector_iter ator<boost::fusion::vector<int, float, float, char, int, int, boost::fusion::void_, boost::fusion::void_, boost::fusion::void_, boost::fusion: :void_>, 0>, boost::fusion::vector_iterator<const boost::fusion::vector<int, float, float, char, int, int, boost::fusion::void_, boost::fusion ::void_, boost::fusion::void_, boost::fusion::void_>, 6> > >]’: problema concepte.cpp:39: instantiated from here /home/thechaos/Escriptori/of_preRelease_v0061_linux_FAT/addons/ofxTableGestures/ext/boost/fusion/functional/invocation/invoke_procedure.hpp:88 : error: incomplete type ‘boost::fusion::detail::invoke_procedure_impl<void (Class2::*)(int, float, float, char, int, int), const boost::fusio n::joint_view<boost::fusion::joint_view<boost::fusion::iterator_range<boost::fusion::vector_iterator<const boost::fusion::vector<int, float, f loat, char, int, int, boost::fusion::void_, boost::fusion::void_, boost::fusion::void_, boost::fusion::void_>, 0>, boost::fusion::vector_itera tor<boost::fusion::vector<int, float, float, char, int, int, boost::fusion::void_, boost::fusion::void_, boost::fusion::void_, boost::fusion:: void_>, 0> >, const boost::fusion::single_view<Class2*> >, boost::fusion::iterator_range<boost::fusion::vector_iterator<boost::fusion::vector< int, float, float, char, int, int, boost::fusion::void_, boost::fusion::void_, boost::fusion::void_, boost::fusion::void_>, 0>, boost::fusion: :vector_iterator<const boost::fusion::vector<int, float, float, char, int, int, boost::fusion::void_, boost::fusion::void_, boost::fusion::voi d_, boost::fusion::void_>, 6> > >, 7, true, false>’ used in nested name specifier However, if I change the number of arguments in the vectors and the method from 6 to 5 from int,float,float,char,int,int to int,float,float,char,int,I can compile it without problems. I suspected about the maximum number of arguments being a limitation, but I tried to change it through defining FUSION_MAX_VECTOR_SIZE without success. I am unable to see what am I doing wrong. Can you reproduce this? Can it be a boost bug (i doubt it but is not impossible)?

  Read the article

• How to configure a WCF service to only accept a single client identified by a x509 certificate

  - by Johan Levin

  I have a WCF client/service app that relies on secure communication between two machines and I want to use use x509 certificates installed in the certificate store to identify the server and client to each other. I do this by configuring the binding as <security authenticationMode="MutualCertificate"/>. There is only client machine. The server has a certificate issued to server.mydomain.com installed in the Local Computer/Personal store and the client has a certificate issued to client.mydomain.com installed in the same place. In addition to this the server has the client's public certificate in Local Computer/Trusted People and the client has the server's public certificate in Local Computer/Trusted People. Finally the client has been configured to check the server's certificate. I did this using the system.servicemodel/behaviors/endpointBehaviors/clientCredentials/serviceCertificate/defaultCertificate element in the config file. So far so good, this all works. My problem is that I want to specify in the server's config file that only clients that identify themselves with the client.mydomain.com certificate from the Trusted People certificate store are allowed to connect. The correct information is available on the server using the ServiceSecurityContext, but I am looking for a way to specify in app.config that WCF should do this check instead of my having to check the security context from code. Is that possible? Any hints would be appreciated. By the way, my server's config file looks like this so far: <?xml version="1.0" encoding="utf-8" ?> <configuration> <system.serviceModel> <services> <service name="MyServer.Server" behaviorConfiguration="CertificateBehavior"> <endpoint contract="Contracts.IMyService" binding="customBinding" bindingConfiguration="SecureConfig"> </endpoint> <host> <baseAddresses> <add baseAddress="http://localhost/SecureWcf"/> </baseAddresses> </host> </service> </services> <behaviors> <serviceBehaviors> <behavior name="CertificateBehavior"> <serviceCredentials> <serviceCertificate storeLocation="LocalMachine" x509FindType="FindBySubjectName" findValue="server.mydomain.com"/> </serviceCredentials> </behavior> </serviceBehaviors> </behaviors> <bindings> <customBinding> <binding name="SecureConfig"> <security authenticationMode="MutualCertificate"/> <httpTransport/> </binding> </customBinding> </bindings> </system.serviceModel> </configuration>

  Read the article

• WCF client encrypt message to JAVA WS using username_token with message protection client policy

  - by Alex

  I am trying to create a WCF client APP that is consuming a JAVA WS that uses username_token with message protection client policy. There is a private key that is installed on the server and a public certificate file was exported from the JKS keystore file. I have installed the public key into certificate store via MMC under Personal certificates. I am trying to create a binding that will encrypt the message and pass the username as part of the payload. I have been researching and trying the different configurations for about a day now. I found a similar situation on the msdn forum: http://social.msdn.microsoft.com/Forums/en/wcf/thread/ce4b1bf5-8357-4e15-beb7-2e71b27d7415 This is the configuration that I am using in my app.config <customBinding> <binding name="certbinding"> <security authenticationMode="UserNameOverTransport"> <secureConversationBootstrap /> </security> <httpsTransport requireClientCertificate="true" /> </binding> </customBinding> <endpoint address="https://localhost:8443/ZZZService?wsdl" binding="customBinding" bindingConfiguration="cbinding" contract="XXX.YYYPortType" name="ServiceEndPointCfg" /> And this is the client code that I am using: EndpointAddress endpointAddress = new EndpointAddress(url + "?wsdl"); P6.WCF.Project.ProjectPortTypeClient proxy = new P6.WCF.Project.ProjectPortTypeClient("ServiceEndPointCfg", endpointAddress); proxy.ClientCredentials.UserName.UserName = UserName; proxy.ClientCredentials.ClientCertificate.SetCertificate(StoreLocation.CurrentUser, StoreName.My, X509FindType.FindByThumbprint, "67 87 ba 28 80 a6 27 f8 01 a6 53 2f 4a 43 3b 47 3e 88 5a c1"); var projects = proxy.ReadProjects(readProjects); This is the .NET CLient error I get: Error Log: Invalid security information. On the Java WS side I trace the log : SEVERE: Encryption is enabled but there is no encrypted key in the request. I traced the SOAP headers and payload and did confirm the encrypted key is not there. Headers: {expect=[100-continue], content-type=[text/xml; charset=utf-8], connection=[Keep-Alive], host=[localhost:8443], Content-Length=[731], vsdebuggercausalitydata=[uIDPo6hC1kng3ehImoceZNpAjXsAAAAAUBpXWdHrtkSTXPWB7oOvGZwi7MLEYUZKuRTz1XkJ3soACQAA], SOAPAction=[""], Content-Type=[text/xml; charset=utf-8]} Payload: <s:Envelope xmlns:s="http://schemas.xmlsoap.org/soap/envelope/" xmlns:u="http://docs.oasis-open.org/wss/2004/01/oasis-200401-wss-wssecurity-utility-1.0.xsd"><s:Header><o:Security s:mustUnderstand="1" xmlns:o="http://docs.oasis-open.org/wss/2004/01/oasis-200401-wss-wssecurity-secext-1.0.xsd"><o:UsernameToken u:Id="uuid-5809743b-d6e1-41a3-bc7c-66eba0a00998-1"><o:Username>admin</o:Username><o:Password>admin</o:Password></o:UsernameToken></o:Security></s:Header><s:Body xmlns:xsi="http://www.w3.org/2001/XMLSchema-instance" xmlns:xsd="http://www.w3.org/2001/XMLSchema"><ReadProjects xmlns="http://xmlns.dev.com/WS/Project/V1"><Field>ObjectId</Field><Filter>Id='WS-Demo'</Filter></ReadProjects></s:Body></s:Envelope> I have also tryed some other bindings but with no success: <basicHttpBinding> <binding name="basicHttp"> <security mode="TransportWithMessageCredential"> <message clientCredentialType="Certificate"/> </security> </binding> </basicHttpBinding> <wsHttpBinding> <binding name="wsBinding"> <security mode="Message"> <message clientCredentialType="UserName" negotiateServiceCredential="false" /> </security> </binding> </wsHttpBinding> Your help will be greatly aprreciatted! Thanks!

  Read the article

• Summer Programming Plans

  - by Gabe

  I've wanted to start "hacking" for many months now. But I put it off in favor of school and other things. Now, though, I'm free for the summer and want to learn as much as I can. I have a rough idea of what I want to try my hand at, but need some guidance as to what specifically - and how - I should learn. This is my plan so far: 1) Get good at programming in general. I plan to read up on how to think/work like a programmer. I'm waiting for the Pragmatic Programmer to arrive, which will be the first book I read. Q: What other books/ebooks should I look at? What more can I do here? 2) Learn/Improve at HTML/CSS. My first project will be to make a personal website/blog for myself using HTML and CSS. ----Then I hope to write/design articles like Dustin Curtis. After I finish this (and learn a programming language) I'll try to create user-based a user-focused website. Q: It's my understanding that just trying to design/manage websites is a good way to learn/improve at HTML/CSS. Is that all correct? 3) Try music development. This might be a sort of stretch for stackoverflow, but I'm interested in mixing/making techno songs. (Think Justice, or Daft Punk, or MSTRKRFT.) Q: I have a Mac. Any ideas on how I could start/learn music making? Any programs I should download, for instance? 4) My main goal: Learning a web development language/framework. I'm a year into learning/using C++. But what I really want to do is develop websites and web apps. I've searched online, and there seems to be great debate over which language/framework to learn first (and which is best). I think I've narrowed it down to three: Ruby (Rails), Python (Django), and PHP (?). Q #1: Which should I learn and use first? (Reasons?) Q #2: One reason I was leaning towards PHP is that I'm taking a PHP development course next semester. Learning it now would make that course easy. If PHP was not the answer to Q #1, is it worth learning both? Or, would it be better to just focus on PHP for this summer and next semester, and then transition thereafter to a better language? 5) iPhone/iPad Programming (Maybe). I've a number of simple, useful app ideas that I'd like to eventually get too. I just bought a Mac, as well as a few app development books. Q #1: Am I spreading myself thin trying to learn all of the above, and objective-C? Q #2: How much harder/easier is objective-C compared to the above languages? Also, how easy is it to learn obj-C after learning a web development language (and some C++)? Q #3: Yes or no? Should I go for it, or just keeep with #1-4 for now? Also: If you have any tips on how I should learn (or how you learned to hack), I'm all ears. I'd be especially interested in how you planned out learning: did you just hack whenever you felt like it, or did you "study" the language a few hours a day, or something else? Thanks so much, guys.

  Read the article

• jQuery dialog box, php form.

  - by tony noriega

  i have a dialog box that opens on pageload for a site. script type="text/javascript"> $(function() { $('#dialog-message').dialog({ modal: 'true', width: '400' }); }); </script> this pulls up an include: <div id="dialog-message" title="Free Jiu Jitsu Session at Alliance"> <!--#include virtual="/includes/guest.php" --> guest.php has a very small form that is processed by the page itself: <?php $dbh=mysql_connect //login stuff here if (isset($_POST['submit'])) { if (!$_POST['name'] | !$_POST['email']) { echo"<div class='error'>Error<br />Please provide your Name and Email Address so we may properly contact you.</div>"; } else { $age = $_POST['age']; $name = $_POST['name']; $gender = $_POST['gender']; $email = $_POST['email']; $phone = $_POST['phone']; $comments = $_POST['comments']; $query = "INSERT INTO table here (age,name,gender,email,phone,comments) VALUES ('$age','$name','$gender','$email','$phone','$comments')"; mysql_query($query); mysql_close(); $yoursite = "my site here"; $youremail = $email; $subject = "Website Guest Contact Us Form"; $message = "message here"; $email2 = "send to email address"; mail($email2, $subject, $message, "From: $email"); echo"<div class='thankyou'>Thank you for contacting us,<br /> we will respond as soon as we can.</div>"; } } ?> <form id="contact_us" class="guest" method="post" action="/guest.php" > <fieldset> <legend>Personal Info</legend> <label for="name" class="guest">Name:</label> <input type="text" name="name" id="name" value="" /><br> <label for="phone" class="guest">Phone:</label> <input type="text" name="phone" id="phone" value="" /><br> <label for="email" class="guest">Email Address:</label> <input type="text" name="email" id="email" value="" /><br> <label for="age" class="guest">Age:</label> <input type="text" name="age" id="age" value="" size="2" /><br> <label for="gender" class="guest">Sex:</label> <input type="radio" name="gender" value="male" /> Male <input type="radio" name="gender" value="female" /> Female<br /> </fieldset> <fieldset> <legend>Comments</legend> <label for="comments" class="guest">Comments / Questions:</label> <textarea id="comments" name="comments" rows="4" cols="22"></textarea><br> <input type="submit" value="Submit" name="submit" /> <input type="Reset" value="Reset" /> </fieldset> </form> Problem is, that the path of the form action does not work, becasue this dialog box is on the index.html page of the site, and if i put the absolute path, it doesnt process... i have this functioning on another contact us page, so i know it works, but wit the dialog box, it seems to have stumped me... what should i do?

  Read the article

• Parsing a Directory of files - Check for a String

  - by i.h4d35

  This is my first post here so kindly pardon any mistakes that I have. I'm still learning to find my way around Stack Exchange. I am trying to write a Java program that tries to scan a Directory full of either .txt,.rtf or.doc files(and none other). The aim is to search all the files in the directory, and find out if a particular string exists in the file. If it does, it returns the string and the filename that it found the string in. The aim of this program is, it is a project for school wherein the program scans the personal folders of call center employees to check if they have stored any CC/DC nos and if yes, report the folder name - to reduce CC fraud. The search function was fairly straight forward and works when I individually specify the filename. However, the searching the directory and passing the files to the search function has me stumped. I've posted my code so far, if you guys could look thru it and give me some feedback/suggestions, I'd really appreciate it. Thanks in advance import java.io.*; import java.util.*; public class parse2{ void traverse(String directory) throws FileNotFoundException { File dir = new File(directory); if (dir.isDirectory()) { String[] children = dir.list(); for (int i=0; i<children.length; i++) { //System.out.println("\n" + children[i]); reader(children[i]); } } } void reader(String loc) throws FileNotFoundException { FileReader fr = new FileReader(loc); BufferedReader br = new BufferedReader(fr); Scanner sc = new Scanner(br); char[] chkArray; int chk=1; char ch; while(sc.hasNext()) { String chkStr = sc.next(); chkArray = chkStr.toCharArray(); if ((chkArray[0]=='4')&&(chkStr.length()>13)) { for(int i=0;i<chkArray.length;i++) { ch=chkArray[i]; if((ch=='0')||(ch=='1')||(ch=='2')||(ch=='3')||(ch=='4')||(ch=='5')||(ch=='6')||(ch=='7')||(ch=='8')||(ch=='9')) { chk=0; continue; } else { chk=1; break; } } if(chk==0) System.out.println("\n"+ chkStr); } else if((chkArray[0]=='5')&&(chkStr.length()>13)) { for(int i=0;i<chkArray.length;i++) { ch=chkArray[i]; if((ch=='0')||(ch=='1')||(ch=='2')||(ch=='3')||(ch=='4')||(ch=='5')||(ch=='6')||(ch=='7')||(ch=='8')||(ch=='9')) { chk=0; continue; } else { chk=1; break; } } if(chk==0) System.out.println("\n"+ chkStr); } else if((chkArray[0]=='6')&&(chkStr.length()>13)) { for(int i=0;i<chkArray.length;i++) { ch=chkArray[i]; if((ch=='0')||(ch=='1')||(ch=='2')||(ch=='3')||(ch=='4')||(ch=='5')||(ch=='6')||(ch=='7')||(ch=='8')||(ch=='9')) { chk=0; continue; } else { chk=1; break; } } if(chk==0) System.out.println("\n"+ chkStr); } } } public static void main(String args[]) throws FileNotFoundException { parse2 P = new parse2(); P.traverse("C:/Documents and Settings/h4d35/Desktop/javatest/chk"); } }

  Read the article

• How do I send an email with embedded images AND regular attachments in JavaMail?

  - by Chris

  Hi, I'd like to know how to build an SMTP multipart message in the correct order so that it will render correctly on the iPhone mail client (rendering correctly in GMail). I'm using Javamail to build up an email containing the following parts: A body part with content type "text/html; UTF-8" An embedded image attachment. A file attachment I am sending the mail via GMail SMTP (via SSL) and the mail is sent and rendered correctly using a GMail account, however, the mail does not render correctly on the iPhone mail client. On the iPhone mail client, the image is rendered before the "Before Image" text when it should be rendered afterwards. After the "Before Image" text there is an icon with a question mark (I assume it means it couldn't find the referenced CID). I'm not sure if this is a limitation of the iPhone mail client or a bug in my mail sending code (I strongly assume the latter). I think that perhaps the headers on my parts might by incorrect or perhaps I am providing the multiparts in the wrong order. I include the text of the received mail as output by gmail (which renders the file correc Message-ID: <[email protected]> Subject: =?UTF-8?Q?Test_from_=E3=82=AF=E3=83=AA=E3=82=B9?= MIME-Version: 1.0 Content-Type: multipart/mixed; boundary="----=_Part_0_20870565.1274154021755" ------=_Part_0_20870565.1274154021755 Content-Type: application/octet-stream Content-Transfer-Encoding: base64 Content-ID: <20100518124021763_368238_0> iVBORw0K ----- TRIMMED FOR CONCISENESS 6p1VVy4alAAAAABJRU5ErkJggg== ------=_Part_0_20870565.1274154021755 Content-Type: text/html; charset=UTF-8 Content-Transfer-Encoding: 7bit <html><head><title>Employees Favourite Foods</title> <style> body { font: normal 8pt arial; } th { font: bold 8pt arial; white-space: nowrap; } td { font: normal 8pt arial; white-space: nowrap; } </style></head><body> Before Image<br><img src="cid:20100518124021763_368238_0"> After Image<br><table border="0"> <tr> <th colspan="4">Employees Favourite Foods</th> </tr> <tr> <th align="left">Name</th><th align="left">Age</th><th align="left">Tel.No</th><th align="left">Fav.Food</th> </tr> <tr style="background-color:#e0e0e0"> <td>Chris</td><td>34</td><td>555-123-4567</td><td>Pancakes</td> </tr> </table></body></html> ------=_Part_0_20870565.1274154021755 Content-Type: text/plain; charset=us-ascii; name=textfile.txt Content-Transfer-Encoding: 7bit Content-Disposition: attachment; filename=textfile.txt This is a textfile with numbers counting from one to ten beneath this line: one two three four five six seven eight nine ten(no trailing carriage return) ------=_Part_0_20870565.1274154021755-- Even if you can't assist me with this, I would appreciate it if any members of the forum could forward me a (non-personal) mail that includes inline images (not external hyperlinked images though). I just need to find a working sample then I can move past this. Thanks, Chris.

  Read the article

• Custom UserControl property not being set via XAML DataBinding in Silverlight 4

  - by programatique

  I have a custom user control called GoalProgressControl. Another user control contains GoalProgressControl and sets its GoalName attribute via databinding in XAML. However, the GoalName property is never set. When I check it in debug mode GoalName remains "null" for the control's lifetime. How do I set the GoalName property? Is there something I am doing incorrectly? I am using .NET Framework 4 and Silverlight 4. I am relatively new to XAML and Silverlight so any help would be greatly appreciated. I have attempted to change GoalProgressControl.GoalName into a POCO property but this causes a Silverlight error, and my reading leads me to believe that databound properties should be of type DependencyProperty. I've also simplified my code to just focus on the GoalName property (the code is below) with no success. Here is GoalProgressControl.xaml: <UserControl x:Class="GoalView.GoalProgressControl" xmlns="http://schemas.microsoft.com/winfx/2006/xaml/presentation" xmlns:x="http://schemas.microsoft.com/winfx/2006/xaml" DataContext="{Binding RelativeSource={RelativeSource Self}}" Height="100"> <Border Margin="5" Padding="5" BorderBrush="#999" BorderThickness="1"> <TextBlock Text="{Binding GoalName}"/> </Border> </UserControl> GoalProgressControl.xaml.cs: public partial class GoalProgressControl : UserControl, INotifyPropertyChanged { public GoalProgressControl() { InitializeComponent(); } public event PropertyChangedEventHandler PropertyChanged; public void NotifyPropertyChanged(string propertyName) { if (PropertyChanged != null) { PropertyChanged(this, new PropertyChangedEventArgs(propertyName)); } } public static DependencyProperty GoalNameProperty = DependencyProperty.Register("GoalName", typeof(string), typeof(GoalProgressControl), null); public string GoalName { get { return (String)GetValue(GoalProgressControl.GoalNameProperty); } set { base.SetValue(GoalProgressControl.GoalNameProperty, value); NotifyPropertyChanged("GoalName"); } } } I've placed GoalProgressControl on another page: <Grid Grid.Row="1" Grid.Column="1" Grid.ColumnSpan="2" Margin="5" Background="#eee" Height="200"> <Border BorderBrush="#999" BorderThickness="1" Background="White"> <StackPanel> <hgc:SectionTitleBar x:Name="ttlGoals" Title="Personal Goals" ImageSource="../Images/check.png" Uri="/Pages/GoalPage.xaml" MoreVisibility="Visible" /> <ItemsControl ItemsSource="{Binding Path=GoalItems}"> <ItemsControl.ItemTemplate> <DataTemplate> <!--TextBlock Text="{Binding Path=[Name]}"/--> <goal:GoalProgressControl GoalName="{Binding Path=[Name]}"/> </DataTemplate> </ItemsControl.ItemTemplate> </ItemsControl> </StackPanel> </Border> </Grid> Please note the commented out "TextBlock" item above. If I comment in the TextBlock and comment out the GoalProgressControl, the binding works correctly and the TextBlock shows the GoalName correctly. Also, if I replace the "GoalName" property above with a simple text string (ex "hello world"), the control renders correctly and "hello world" is shown on the control when it renders.

  Read the article

• A methology that allows for a single Java code base covering many different versions?

  - by Thorbjørn Ravn Andersen

  I work in a small shop where we have a LOT of legacy Cobol code and where a methology has been adopted to allow us to minimize forking and branching as much as possible. For a given release we have three levels: CORE - bottom layer, this code is common to all releases GROUP - optional code common to several customers. CUSTOMER - optional code specific for a single customer. When a program is needed, it is first searched for in CUSTOMER, then in GROUP and finally in CORE. A given application for us invokes many programs which all are looked for in this sequence (think exe files and PATH under Windows). We also have Java programs interacting with this legacy code, and as the core-group-customer lookup mehchanism does not lend it self easily to Java it has tended to grow in a CVS branch for each customer, requiring much too much maintainance. The Java part and the backend part tend to be developed in parallel. I have been assigned to figure out a way to make the two worlds meet. Essentially we want a Java enviornment which allows us to have a single code base with sources for each release, where we easily can select a group and a customer and work with the application as it goes for that customer, and then easily switch to another codeset and THAT customer. I was thinking of perhaps a scenario with an Eclipse project for each core, customer, and group and then use Project Sets to select those we need for a given scenario. The problem I cannot get my head about, is how we would create robust code in the CORE projects which will work regardless of which group and customer is selected. A Factory class which knows which sub class of a passed Class object to invoke instead of each and every new? Others must have had similar code base management problems. Anybody with experiences to share? EDIT: The conclusion to this problem above has been that CVS needs to be replaced with a source code management system better suited for dealing with many branches concurrently and the migration of source from one component to the other while keeping history. Inspired by the recent migration by slf4j and logback we are currently looking at git as it handles branches very well. We've considered subversion and mercurial too but git appears to be better for single location, multibranched projects. I've asked about Perforce in another question, but my personal inclination is towards open source solutions for something as crucial as this. EDIT: After some more pondering, we've found that our actual pain point is that we use branches in CVS, and that branches in CVS are the easiest to work with if you branch ALL files! The revised conclusion is that we can do this with CVS alone, by switching to a forest of java projects, each corresponding to one of the levels above, and use the Eclipse build paths to tie them together so each CUSTOMER version pulls in the appropriate GROUP and CORE project. We still want to switch to a better versioning system but this is so important a decision so we want to delay it as much as possible. EDIT: I now have a proof-of-concept implementation of the CORE-GROUP-CUSTOMER concept using Google Guice 2.0 - the @ImplementedBy tag is just what we need. I wonder what everybody else does? Using if's all over the place? EDIT: Now I also need this functionality for web applications. Guice was until the JSR-330 is in place. Anybody with versioning experience? EDIT: JSR-330/299 is now in place with the JEE6 reference implementation Weld based on JBoss Seam and I have reimplemented the proof-of-concept with Weld and can see that if we use @Alternative along with ... in beans.xml we can get the behaviour we desire. I.e. provide a new implementation for a given functionality in CORE without changing a bit in the CORE jars. Initial reading up on the Servlet 3.0 specification indicates that it may support the same functionality for web application resources (not code). We will now do initial testing on the real application.

  Read the article

• What is "elegant" code?

  - by Breton

  I see a lot of lip service and talk about the most "elegant" way to do this or that. I think if you spend enough time programming you begin to obtain a sort of intuitive feel for what it is we call "elegance". But I'm curious. Even if we can look at a bit of code, and say instinctively "That's elegant", or "That's messy", I wonder if any of us really understands what that means. Is there a precise definition for this "elegance" we keep referring to? If there is, what is it? Now, what I mean by a precise definition, is a series of statements which can be used to derive questions about a peice of code, or a program as a whole, and determine objectively, or as objectively as possible, whether that code is "elegant" or not. May I assert, that perhaps no such definition exists, and it's all just personal preference. In this case, I ask you a slightly different question: Is there a better word for "elegance", or a better set of attributes to use for judging code quality that is perhaps more objective than merely appealing to individual intuition and taste? Perhaps code quality is a matter of taste, and the answer to both of my questions is "no". But I can't help but feel that we could be doing better than just expressing wishy washy feelings about our code quality. For example, user interface design is something that to a broad range of people looks for all the world like a field of study that oughtta be 100% subjective matter of taste. But this is shockingly and brutally not the case, and there are in fact many objective measures that can be applied to a user interface to determine its quality. A series of tests could be written to give a definitive and repeatable score to user interface quality. (See GOMS, for instance). Now, okay. is Elegance simply "code quality" or is it something more? Is it something that can be measured? Or is it a matter of taste? Does our profession have room for taste? Maybe I'm asking the wrong questions altogether. Help me out here. Bonus Round If there is such a thing as elegance in code, and that concept is useful, do you think that justifies classifying the field of programming as an "Art" capital A, or merely a "craft". Or is it just an engineering field populated by a bunch of wishful thinking humans? Consider this question in the light of your thoughts about the elegance question. Please note that there is a distinction between code which is considered "art" in itself, and code that was written merely in the service of creating an artful program. When I ask this question, I ask if the code itself justifies calling programming an art. Bounty Note I liked the answers to this question so much, I think I'd like to make a photographic essay book from it. Released as a free PDF, and published on some kind of on demand printing service of course, such as "zazz" or "tiggle" or "printley" or something . I'd like some more answers, please!

  Read the article

• Order of parts in SMTP multipart messages

  - by Chris

  Hi, I'd like to know how to build an SMTP multipart message in the correct order so that it will render correctly on the iPhone mail client (rendering correctly in GMail). I'm using Javamail to build up an email containing the following parts: A body part with content type "text/html; UTF-8" An embedded image attachment. A file attachment I am sending the mail via GMail SMTP (via SSL) and the mail is sent and rendered correctly using a GMail account, however, the mail does not render correctly on the iPhone mail client. On the iPhone mail client, the image is rendered before the "Before Image" text when it should be rendered afterwards. After the "Before Image" text there is an icon with a question mark (I assume it means it couldn't find the referenced CID). I'm not sure if this is a limitation of the iPhone mail client or a bug in my mail sending code (I strongly assume the latter). I think that perhaps the headers on my parts might by incorrect or perhaps I am providing the multiparts in the wrong order. I include the text of the received mail as output by gmail (which renders the file correc Message-ID: <[email protected]> Subject: =?UTF-8?Q?Test_from_=E3=82=AF=E3=83=AA=E3=82=B9?= MIME-Version: 1.0 Content-Type: multipart/mixed; boundary="----=_Part_0_20870565.1274154021755" ------=_Part_0_20870565.1274154021755 Content-Type: application/octet-stream Content-Transfer-Encoding: base64 Content-ID: <20100518124021763_368238_0> iVBORw0K ----- TRIMMED FOR CONCISENESS 6p1VVy4alAAAAABJRU5ErkJggg== ------=_Part_0_20870565.1274154021755 Content-Type: text/html; charset=UTF-8 Content-Transfer-Encoding: 7bit <html><head><title>Employees Favourite Foods</title> <style> body { font: normal 8pt arial; } th { font: bold 8pt arial; white-space: nowrap; } td { font: normal 8pt arial; white-space: nowrap; } </style></head><body> Before Image<br><img src="cid:20100518124021763_368238_0"> After Image<br><table border="0"> <tr> <th colspan="4">Employees Favourite Foods</th> </tr> <tr> <th align="left">Name</th><th align="left">Age</th><th align="left">Tel.No</th><th align="left">Fav.Food</th> </tr> <tr style="background-color:#e0e0e0"> <td>Chris</td><td>34</td><td>555-123-4567</td><td>Pancakes</td> </tr> </table></body></html> ------=_Part_0_20870565.1274154021755 Content-Type: text/plain; charset=us-ascii; name=textfile.txt Content-Transfer-Encoding: 7bit Content-Disposition: attachment; filename=textfile.txt This is a textfile with numbers counting from one to ten beneath this line: one two three four five six seven eight nine ten(no trailing carriage return) ------=_Part_0_20870565.1274154021755-- Even if you can't assist me with this, I would appreciate it if any members of the forum could forward me a (non-personal) mail that includes inline images (not external hyperlinked images though). I just need to find a working sample then I can move past this. Thanks, Chris.

  Read the article

• UIView, UIScrollView and UITextFields problem calling Method

  - by Jeff Groby

  I have a view with several embedded UITextFields, this UIView is subordinated to a UIScrollView in IB. Each text field is supposed to invoke a method called updateText defined in the viewcontroller implementation file when the user is done editing the field. For some reason, the method updateText never gets invoked. Anyone have any ideas how to go about fixing this? The method fired off just fine when the UIScrollView was not present in the project but the keyboard would cover the text fields during input, which was annoying. Now my textfields move up above the keyboard when it appears, but won't fire off the method when done editing. Here is my implementation file: #import "MileMarkerViewController.h" @implementation MileMarkerViewController @synthesize scrollView,milemarkerLogDate,milemarkerDesc,milemarkerOdobeg,milemarkerOdoend,milemarkerBusiness,milemarkerPersonal,milemarker; - (id)initWithNibName:(NSString *)nibNameOrNil bundle:(NSBundle *)nibBundleOrNil { if (self = [super initWithNibName:nibNameOrNil bundle:nibBundleOrNil]) { // Initialization code } return self; } - (BOOL) textFieldShouldReturn: (UITextField*) theTextField { return [theTextField resignFirstResponder]; } - (void)viewDidLoad { [super viewDidLoad]; [[NSNotificationCenter defaultCenter] addObserver: self selector: @selector(keyboardWasShown:) name: UIKeyboardDidShowNotification object: nil]; [[NSNotificationCenter defaultCenter] addObserver: self selector: @selector(keyboardWasHidden:) name: UIKeyboardDidHideNotification object: nil]; keyboardShown = NO; // 1 [scrollView setContentSize: CGSizeMake( 320, 480)]; // 2 } - (void)keyboardWasShown:(NSNotification*)aNotification { if (keyboardShown) return; NSDictionary* info = [aNotification userInfo]; // Get the size of the keyboard. NSValue* aValue = [info objectForKey:UIKeyboardBoundsUserInfoKey]; CGSize keyboardSize = [aValue CGRectValue].size; // Resize the scroll view (which is the root view of the window) CGRect viewFrame = [scrollView frame]; viewFrame.size.height -= keyboardSize.height; scrollView.frame = viewFrame; // Scroll the active text field into view. CGRect textFieldRect = [activeField frame]; [scrollView scrollRectToVisible:textFieldRect animated:YES]; keyboardShown = YES; } - (void)keyboardWasHidden:(NSNotification*)aNotification { NSDictionary* info = [aNotification userInfo]; // Get the size of the keyboard. NSValue* aValue = [info objectForKey:UIKeyboardBoundsUserInfoKey]; CGSize keyboardSize = [aValue CGRectValue].size; // Reset the height of the scroll view to its original value CGRect viewFrame = [scrollView frame]; viewFrame.size.height += keyboardSize.height; [scrollView scrollRectToVisible:CGRectMake(0, 0, 1, 1) animated:YES]; scrollView.frame = viewFrame; keyboardShown = NO; } - (void)textFieldDidBeginEditing:(UITextField *)textField { activeField = textField; } - (void)textFieldDidEndEditing:(UITextField *)textField { activeField = nil; } - (IBAction)updateText:(id) sender { NSLog(@"You just entered: %@",self.milemarkerLogDate.text); self.milemarker.logdate = self.milemarkerLogDate.text; self.milemarker.desc = self.milemarkerDesc.text; self.milemarker.odobeg = self.milemarkerOdobeg.text; self.milemarker.odoend = self.milemarkerOdoend.text; self.milemarker.business = self.milemarkerBusiness.text; self.milemarker.personal = self.milemarkerPersonal.text; NSLog(@"Original textfield is set to: %@",self.milemarker.logdate); [self.milemarker updateText]; } - (void)didReceiveMemoryWarning { // Releases the view if it doesn't have a superview. [super didReceiveMemoryWarning]; // Release any cached data, images, etc that aren't in use. } - (void)dealloc { [super dealloc]; } @end

  Read the article

• Why does this valid Tkinter code crash when mixed with a bit of PyWin32?

  - by Erlog

  So I'm making a very small program for personal use in tkinter, and I've run into a really strange wall. I'm mixing tkinter with the pywin32 bindings because I really hate everything to do with the syntax and naming conventions of pywin32, and it feels like tkinter gets more done with far less code. The strangeness is happening in the transition between the pywin32 clipboard watching and my program's reaction to it in tkinter. My window and all its controls are being handled in tkinter. The pywin32 bindings are doing clipboard watching and clipboard access when the clipboard changes. From what I've gathered about the way the clipboard watching pieces of pywin32 work, you can make it work with anything you want as long as you provide pywin32 with the hwnd value of your window. I'm doing that part, and it works when the program first starts. It just doesn't seem to work when the clipboard changes. When the program launches, it grabs the clipboard and puts it into the search box and edit box just fine. When the clipboard is modified, the event I want to fire off is firing off...except that event that totally worked before when the program launched is now causing a weird hang instead of doing what it's supposed to do. I can print the clipboard contents to stdout all I want if the clipboard changes, but not put that same data into a tkinter widget. It only hangs like that if it starts to interact with any of my tkinter widgets after being fired off by a clipboard change notification. It feels like there's some pywin32 etiquette I've missed in adapting the clipboard-watching sample code I was using over to my tkinter-using program. Tkinter apparently doesn't like to produce stack traces or error messages, and I can't really even begin to know what to look for trying to debug it with pdb. Here's the code: #coding: utf-8 #Clipboard watching cribbed from ## {{{ http://code.activestate.com/recipes/355593/ (r1) import pdb from Tkinter import * import win32clipboard import win32api import win32gui import win32con import win32clipboard def force_unicode(object, encoding="utf-8"): if isinstance(object, basestring) and not isinstance(object, unicode): object = unicode(object, encoding) return object class Application(Frame): def __init__(self, master=None): self.master = master Frame.__init__(self, master) self.pack() self.createWidgets() self.hwnd = self.winfo_id() self.nextWnd = None self.first = True self.oldWndProc = win32gui.SetWindowLong(self.hwnd, win32con.GWL_WNDPROC, self.MyWndProc) try: self.nextWnd = win32clipboard.SetClipboardViewer(self.hwnd) except win32api.error: if win32api.GetLastError () == 0: # information that there is no other window in chain pass else: raise self.update_search_box() self.word_search() def word_search(self): #pdb.set_trace() term = self.searchbox.get() self.resultsbox.insert(END, term) def update_search_box(self): clipboardtext = "" if win32clipboard.IsClipboardFormatAvailable(win32clipboard.CF_TEXT): win32clipboard.OpenClipboard() clipboardtext = win32clipboard.GetClipboardData() win32clipboard.CloseClipboard() if clipboardtext != "": self.searchbox.delete(0,END) clipboardtext = force_unicode(clipboardtext) self.searchbox.insert(0, clipboardtext) def createWidgets(self): self.button = Button(self) self.button["text"] = "Search" self.button["command"] = self.word_search self.searchbox = Entry(self) self.resultsbox = Text(self) #Pack everything down here for "easy" layout changes later self.searchbox.pack() self.button.pack() self.resultsbox.pack() def MyWndProc (self, hWnd, msg, wParam, lParam): if msg == win32con.WM_CHANGECBCHAIN: self.OnChangeCBChain(msg, wParam, lParam) elif msg == win32con.WM_DRAWCLIPBOARD: self.OnDrawClipboard(msg, wParam, lParam) # Restore the old WndProc. Notice the use of win32api # instead of win32gui here. This is to avoid an error due to # not passing a callable object. if msg == win32con.WM_DESTROY: if self.nextWnd: win32clipboard.ChangeClipboardChain (self.hwnd, self.nextWnd) else: win32clipboard.ChangeClipboardChain (self.hwnd, 0) win32api.SetWindowLong(self.hwnd, win32con.GWL_WNDPROC, self.oldWndProc) # Pass all messages (in this case, yours may be different) on # to the original WndProc return win32gui.CallWindowProc(self.oldWndProc, hWnd, msg, wParam, lParam) def OnChangeCBChain (self, msg, wParam, lParam): if self.nextWnd == wParam: # repair the chain self.nextWnd = lParam if self.nextWnd: # pass the message to the next window in chain win32api.SendMessage (self.nextWnd, msg, wParam, lParam) def OnDrawClipboard (self, msg, wParam, lParam): if self.first: self.first = False else: #print "changed" self.word_search() #self.word_search() if self.nextWnd: # pass the message to the next window in chain win32api.SendMessage(self.nextWnd, msg, wParam, lParam) if __name__ == "__main__": root = Tk() app = Application(master=root) app.mainloop() root.destroy()

  Read the article

• evaluating cost/benefits of using extension methods in C# => 3.0

  - by BillW

  Hi, In what circumstances (usage scenarios) would you choose to write an extension rather than sub-classing an object ? < full disclosure : I am not an MS employee; I do not know Mitsu Furota personally; I do know the author of the open-source Componax library mentioned here, but I have no business dealings with him whatsoever; I am not creating, or planning to create any commercial product using extensions : in sum : this post is from pure intellectal curiousity related to my trying to (continually) become aware of "best practices" I find the idea of extension methods "cool," and obviously you can do "far-out" things with them as in the many examples you can in Mitsu Furota's (MS) blog postslink text. A personal friend wrote the open-source Componax librarylink text, and there's some remarkable facilities in there; but he is in complete command of his small company with total control over code guidelines, and every line of code "passes through his hands." While this is speculation on my part : I think/guess other issues might come into play in a medium-to-large software team situation re use of Extensions. Looking at MS's guidelines at link text, you find : In general, you will probably be calling extension methods far more often than implementing your own. ... In general, we recommend that you implement extension methods sparingly and only when you have to. Whenever possible, client code that must extend an existing type should do so by creating a new type derived from the existing type. For more information, see Inheritance (C# Programming Guide). ... When the compiler encounters a method invocation, it first looks for a match in the type's instance methods. If no match is found, it will search for any extension methods that are defined for the type, and bind to the first extension method that it finds. And at Ms's link text : Extension methods present no specific security vulnerabilities. They can never be used to impersonate existing methods on a type, because all name collisions are resolved in favor of the instance or static method defined by the type itself. Extension methods cannot access any private data in the extended class. Factors that seem obvious to me would include : I assume you would not write an extension unless you expected it be used very generally and very frequently. On the other hand : couldn't you say the same thing about sub-classing ? Knowing we can compile them into a seperate dll, and add the compiled dll, and reference it, and then use the extensions : is "cool," but does that "balance out" the cost inherent in the compiler first having to check to see if instance methods are defined as described above. Or the cost, in case of a "name clash," of using the Static invocation methods to make sure your extension is invoked rather than the instance definition ? How frequent use of Extensions would affect run-time performance or memory use : I have no idea. So, I'd appreciate your thoughts, or knowing about how/when you do, or don't do, use Extensions, compared to sub-classing. thanks, Bill

  Read the article

• C++ Compile problem when using Windows - CodeGear

  - by Carlos

  This is a follow-up question to this one i made earlier. Btw thanks Neil Butterworth for you help http://stackoverflow.com/questions/2461977/problem-compiling-c-in-codegear A quick recap. Im currently developing a C++ program for university, I used Netbeans 6.8 on my personal computer (Mac) and all works perfect. When I try them on my windows partition or at the university PC's using CodeGear RAD Studio 2009 & 2010 i was getting a few compile errors which were solved by adding the following header file: #include <string> However now the program does compile but it doesn't run, just a blank console. And am getting the following in the CodeGear event's log: Thread Start: Thread ID: 2024. Process Project1.exe (3280) Process Start: C:\Users\Carlos\Documents\RAD Studio\Projects\Debug\Project1.exe. Base Address: $00400000. Process Project1.exe (3280) Module Load: Project1.exe. Has Debug Info. Base Address: $00400000. Process Project1.exe (3280) Module Load: ntdll.dll. No Debug Info. Base Address: $77E80000. Process Project1.exe (3280) Module Load: KERNEL32.dll. No Debug Info. Base Address: $771C0000. Process Project1.exe (3280) Module Load: KERNELBASE.dll. No Debug Info. Base Address: $75FE0000. Process Project1.exe (3280) Module Load: cc32100.dll. No Debug Info. Base Address: $32A00000. Process Project1.exe (3280) Module Load: USER32.dll. No Debug Info. Base Address: $77980000. Process Project1.exe (3280) Module Load: GDI32.dll. No Debug Info. Base Address: $75F50000. Process Project1.exe (3280) Module Load: LPK.dll. No Debug Info. Base Address: $75AB0000. Process Project1.exe (3280) Module Load: USP10.dll. No Debug Info. Base Address: $76030000. Process Project1.exe (3280) Module Load: msvcrt.dll. No Debug Info. Base Address: $776A0000. Process Project1.exe (3280) Module Load: ADVAPI32.dll. No Debug Info. Base Address: $777D0000. Process Project1.exe (3280) Module Load: SECHOST.dll. No Debug Info. Base Address: $77960000. Process Project1.exe (3280) Module Load: RPCRT4.dll. No Debug Info. Base Address: $762F0000. Process Project1.exe (3280) Module Load: SspiCli.dll. No Debug Info. Base Address: $759F0000. Process Project1.exe (3280) Module Load: CRYPTBASE.dll. No Debug Info. Base Address: $759E0000. Process Project1.exe (3280) Module Load: IMM32.dll. No Debug Info. Base Address: $763F0000. Process Project1.exe (3280) Module Load: MSCTF.dll. No Debug Info. Base Address: $75AD0000. Process Project1.exe (3280) I would really appreciate any help or ideas on how to solve this problem. P.S: In the case anyone wonders why am I sticking with CodeGear is because is the IDE professors use to evaluate our assignments.

  Read the article

• ASP.NET gridview control in side update panel has a problem

  - by Eyla

  Greetings, I have gridview with SelectedIndexChanged event. when I click on a record in gridview it should call the SelectedIndexChanged event and do some operations. SelectedIndexChanged event is working OK, but when I put the gridview inside ajax updatepanle SelectedIndexChanged event will not response even if I add AsyncPostBackTrigger trigger for SelectedIndexChanged event. Please look at my code and advice me what I should do!! Thank you <%@ Page Title="" Language="C#" MasterPageFile="~/Master.Master" AutoEventWireup="true" CodeBehind="WebForm1.aspx.cs" Inherits="IMAM_APPLICATION.WebForm1" %> <%@ Register Assembly="AjaxControlToolkit" Namespace="AjaxControlToolkit" TagPrefix="cc1" %> <asp:Content ID="Content1" ContentPlaceHolderID="ContentPlaceHolder1" runat="server"> <div id="mydiv"> <asp:UpdatePanel ID="UpdatePanel1" runat="server"> <ContentTemplate> <br /> <br /> <br /> <asp:GridView ID="GridView1" runat="server" style="position:absolute; top: 280px; left: 30px; height: 240px; width: 915px;" PageSize="5" onselectedindexchanged="GridView1_SelectedIndexChanged" AutoGenerateColumns="False" DataKeyNames="idcontact_info"> <Columns> <asp:CommandField ShowSelectButton="True" /> <asp:BoundField AccessibleHeaderText="Midle Name" DataField="Midle_Name" /> <asp:BoundField DataField="Last_Name" HeaderText="Last Name" /> <asp:BoundField DataField="Phone_home" HeaderText="Phone Home" /> <asp:BoundField DataField="cell_home" HeaderText="Mobile Home" /> <asp:BoundField DataField="phone_work" HeaderText="Phone Work" /> <asp:BoundField DataField="cell_Work" HeaderText="Mobile Work" /> <asp:BoundField DataField="Email_Home" HeaderText="Personal Home" /> <asp:BoundField DataField="Email_work" HeaderText="Work Email" /> </Columns> </asp:GridView> <br /> <br /> <br /></ContentTemplate> <Triggers> <asp:AsyncPostBackTrigger ControlID="GridView1" EventName="SelectedIndexChanged" /> </Triggers> </asp:UpdatePanel> <asp:RadioButton ID="rdoSearchFlat" runat="server" style="position:absolute; top: 565px; left: 70px;" Text="Flat Search" GroupName="Search"/> <asp:TextBox ID="txtSearch" runat="server" style="position:absolute; top: 560px; left: 170px;" ></asp:TextBox> <asp:Button ID="btnSearch" runat="server" Text="Search" style="position:absolute; top: 555px; left: 375px;" CausesValidation="False" onclick="btnSearch_Click"/> <asp:Label ID="Label7" runat="server" Style="position: absolute; top: 630px; left: 85px;" Text="First Name"></asp:Label> <asp:TextBox ID="txtFirstName" runat="server" Style="top: 630px; left: 185px; position: absolute; height: 22px; width: 128px"> </asp:TextBox> </div> </asp:Content>

  Read the article

• Better Way To Use C++ Named Parameter Idiom?

  - by Head Geek

  I've been developing a GUI library for Windows (as a personal side project, no aspirations of usefulness). For my main window class, I've set up a hierarchy of option classes (using the Named Parameter Idiom), because some options are shared and others are specific to particular types of windows (like dialogs). The way the Named Parameter Idiom works, the functions of the parameter class have to return the object they're called on. The problem is that, in the hierarchy, each one has to be a different class -- the createWindowOpts class for standard windows, the createDialogOpts class for dialogs, and the like. I've dealt with that by making all the option classes templates. Here's an example: template <class T> class _sharedWindowOpts: public detail::_baseCreateWindowOpts { public: /////////////////////////////////////////////////////////////// // No required parameters in this case. _sharedWindowOpts() { }; typedef T optType; // Commonly used options optType& at(int x, int y) { mX=x; mY=y; return static_cast<optType&>(*this); }; // Where to put the upper-left corner of the window; if not specified, the system sets it to a default position optType& at(int x, int y, int width, int height) { mX=x; mY=y; mWidth=width; mHeight=height; return static_cast<optType&>(*this); }; // Sets the position and size of the window in a single call optType& background(HBRUSH b) { mBackground=b; return static_cast<optType&>(*this); }; // Sets the default background to this brush optType& background(INT_PTR b) { mBackground=HBRUSH(b+1); return static_cast<optType&>(*this); }; // Sets the default background to one of the COLOR_* colors; defaults to COLOR_WINDOW optType& cursor(HCURSOR c) { mCursor=c; return static_cast<optType&>(*this); }; // Sets the default mouse cursor for this window; defaults to the standard arrow optType& hidden() { mStyle&=~WS_VISIBLE; return static_cast<optType&>(*this); }; // Windows are visible by default optType& icon(HICON iconLarge, HICON iconSmall=0) { mIcon=iconLarge; mSmallIcon=iconSmall; return static_cast<optType&>(*this); }; // Specifies the icon, and optionally a small icon // ...Many others removed... }; template <class T> class _createWindowOpts: public _sharedWindowOpts<T> { public: /////////////////////////////////////////////////////////////// _createWindowOpts() { }; // These can't be used with child windows, or aren't needed optType& menu(HMENU m) { mMenuOrId=m; return static_cast<optType&>(*this); }; // Gives the window a menu optType& owner(HWND hwnd) { mParentOrOwner=hwnd; return static_cast<optType&>(*this); }; // Sets the optional parent/owner }; class createWindowOpts: public _createWindowOpts<createWindowOpts> { public: /////////////////////////////////////////////////////////////// createWindowOpts() { }; }; It works, but as you can see, it requires a noticeable amount of extra work: a type-cast on the return type for each function, extra template classes, etcetera. My question is, is there an easier way to implement the Named Parameter Idiom in this case, one that doesn't require all the extra stuff?

  Read the article

< Previous Page | 363 364 365 366 367 368 369 370 371 372 373 374  | Next Page >