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  • how to pass vector of string to foo(char const *const *const)?

    - by user347208
    Hi, This is my first post so please be nice. I searched in this forum and googled but I still can not find the answer. This problem has bothered me for more than a day, so please give me some help. Thank you. I need to pass a vector of string to a library function foo(char const *const *const). I can not pass the &Vec[0] since it's a pointer to a string. Therefore, I have an array and pass the c_str() to that array. The following is my code (aNames is the vector of string): const char* aR[aNames.size()]; std::transform(aNames.begin(), aNames.end(), aR, boost::bind(&std::string::c_str, _1)); foo(aR); However, it seems it causes some undefined behavior: If I run the above code, then the function foo throw some warnings about illegal characters ('èI' blablabla) in aR. If I print aR before function foo like this: std::copy(aR, aR+rowNames.size(), std::ostream_iterator<const char*>(std::cout, "\n")); foo(aR); Then, everything is fine. My questions are: Does the conversion causes undefined behavior? If so, why? What is the correct way to pass vector of string to foo(char const *const *const)? Thank you very much for your help!

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  • Connectivity issues with dual NIC machine in EC2

    - by Matt Sieker
    I'm trying to get some servers set up in EC2 in a Virtual Private Cloud. To do this, I have two subnets: 10.0.42.0/24 - Public subnet 10.0.83.0/24 - Private subnet To bridge these two, I have a Funtoo instance with a pair of NICs: eth0 10.0.42.10 eth1 10.0.83.10 Which has the following routing table: Kernel IP routing table Destination Gateway Genmask Flags Metric Ref Use Iface 10.0.83.0 * 255.255.255.0 U 0 0 0 eth1 10.0.83.0 * 255.255.255.0 U 203 0 0 eth1 10.0.42.0 * 255.255.255.0 U 202 0 0 eth0 loopback * 255.0.0.0 U 0 0 0 lo default 10.0.42.1 0.0.0.0 UG 0 0 0 eth0 default 10.0.42.1 0.0.0.0 UG 202 0 0 eth0 An elastic IP is attached to the eth0 interface, and I can connect to it fine remotely. However, I cannot ping anything in the 10.0.83.0 subnet. For now iptables is not set up on the box, so there's no rules that would get in the way (Eventually this will be managed by Shorewall, but I should get basic connectivity done first) Subnet details from the VPC interface: CIDR: 10.0.83.0/24 Destination Target 10.0.0.0/16 local 0.0.0.0/0 [ID of eth1 on NAT box] Network ACL: Default Inbound: Rule # Port (Service) Protocol Source Allow/Deny 100 ALL ALL 0.0.0.0/0 ALLOW * ALL ALL 0.0.0.0/0 DENY Outbound: Rule # Port (Service) Protocol Destination Allow/Deny 100 ALL ALL 0.0.0.0/0 ALLOW * ALL ALL 0.0.0.0/0 DENY   CIDR: 10.0.83.0/24 VPC: Destination Target 10.0.0.0/16 local 0.0.0.0/0 [Internet Gateway ID] Network ACL: Default (replace) Inbound: Rule # Port (Service) Protocol Source Allow/Deny 100 ALL ALL 0.0.0.0/0 ALLOW * ALL ALL 0.0.0.0/0 DENY Outbound: Rule # Port (Service) Protocol Destination Allow/Deny 100 ALL ALL 0.0.0.0/0 ALLOW * ALL ALL 0.0.0.0/0 DENY I've been trying to work this out most of the evening, but I'm just stuck. I'm either missing something obvious, or am doing something very wrong. I would think I'd be able to ping from either interface on this box without issue. Hopefully some more pairs of eyes on this configuration will help. EDIT: I am an idiot. After I bothered to install nmap to run some more tests, I discover I can see the ports, and connect to them, pings are just being blocked.

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  • How do I combine static and dynamic DHCP leases on a Cisco router?

    - by Brad
    Basically, what I need is super similar to the unanswered cisco forum question below: https://supportforums.cisco.com/message/3139749#3139749 I have a Cisco 850 Series router. I have configured a DHCP pool for the 10.0.0.0/24 network. I have excluded 10.0.0.1 - 10.0.0.99 from the DHCP pool. I want to add a static DHCP pool for stuff and I want DHCP to statically assign them the addresses of my choice below 100. Actually, I don't care what addresses I statically assign. They can be anything in the pool for all I care, I just want it to work. Why are you doing this? Just statically assign the IPs on the devices! I don't want to do this because I have some laptop users. They could obviously only use that static IP here. This isn't a problem if they could be bothered to change any location setting or something. They can't. So it HAS to be DHCP. It also has to be static IPs because I need to forward ports to them. I know, I know, this is weird but it's an apartment LAN/WLAN so this isn't exactly a typical use case. Relevant sections of config below: ip dhcp excluded-address 10.0.0.1 10.0.0.99 ! ip dhcp pool Internal-net import all network 10.0.0.0 255.255.255.0 default-router 10.0.0.1 domain-name 1770.local lease 7 ! ip dhcp pool static-pool import all origin file flash://staticmap default-router 10.0.0.1 domain-name 1770.local Contents of staticmap: *time* Aug 5 2010 09:00 AM *version* 2 !IP address Type Hardware address Lease expiration 10.0.0.100/24 1 001f.5b3e.d50a Infinite *end* You can see here I was trying addresses outside the excluded-address range to see if that would make any difference. My testing machine's MAC: mainframe:~ brad$ ifconfig en1 en1: flags=8863<UP,BROADCAST,SMART,RUNNING,SIMPLEX,MULTICAST> mtu 1500 ether 00:1f:5b:3e:d5:0a What shows up in the DHCP binding table: basestar#show ip dhcp binding Bindings from all pools not associated with VRF: IP address Client-ID/ Lease expiration Type Hardware address/ User name 10.0.0.112 0100.1f5b.3ed5.0a Aug 12 2010 10:06 AM Automatic What's up with the funny looking MAC in the DHCP binding table?? Is what I'm trying to accomplish basically impossible? Am I going about this the wrong way? All I want to to be able to port forward some ports to specific devices. The way I would do this with a consumer router is to do what I'm trying to do here; assign static DHCP to those devices then configure PAT for ports on those addresses.

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  • In search of a good audio player for Ubuntu 9.10

    - by Joe Casadonte
    If this should be marked Community Wiki, please let me know. I'm switching from XP to Ubuntu, and I have been very disappointed with the selection of media players available. I'm primarily interested in an audio player, but integrated video and library management is OK, too. My criteria: Must be able to play audio CDs (I'm shocked how many apps this does away with, right away) Must be able to play MP3 & WAV; OGG, SHN, FLAC are all bonuses Repeat and Shuffle modes are a must FreeDB / GraceNote through a proxy is a must (if it can read a PAC file, that would be awesome) It needs to be really small, e.g. skinnable or an applet Ability to execute a playlist is a plus Gapless MP3 playback a plus I'm running Gnome, but I'm not totally adverse to a KDE app. Command-line only is also a viable option. Some that I've tried: RhythmBox - probably the best of the lot that I've tried; I don't like its mini mode (doesn't show the song being played) and I can't figure out how to get it to hit FreeDB/GraceNote through a proxy Songbird - can't play CDs, playlist management is atrocious Banshee Jajuk Maybe a couple of more. Thanks! UPDATE I tried out VLC, Amarok and Songbord (again). VLC I eventually got to work (I had some kind of bad configuration). It seemed way more involved than I was looking for out of a music player, and in general more geared to video than audio. I couldn't fathom its library management, which I think it has; maybe it doesn't, and that's why I couldn't figure it out. Amaork looked very promising but the library management was not to my liking, and the way it handled a playlist with both MP3 and WAV is inexplicable at best. I did like some aspects of the UI, but not enough to keep it. Songbird is very finicky, but I like the library management. Sort of. It kept telling me my Watch folder was invalid, even thought it clearly was accessible. Playlist management is bizarre, and the message that it was deleting source files whenever I deleted a playlist had me too worried to keep using it. Had it been able to play CDs, maybe I would have persevered. Audacious, while a bit odd at times, does seem to do what I want. If it had a library manager, I wouldn't have bothered trying any of the others. Thanks for the help, everyone!

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  • What the best way to achieve RPO of zero and lowest possible RTO (less than 15 minutes) with SQL 2008 R2?

    - by Adrian Hope-Bailie
    We are running a payments (EFT transaction processing) application which is processing high volumes of transactions 24/7 and are currently investigating a better way of doing DB replication to our disaster recovery site. Our current and previous strategies have included using both DoubleTake and Redgate to replicate data to a warm stand-by. DoubleTake is the supported solution from the payments software vendor however their (DoubleTake's) support in South Africa is very poor. We had a few issues and simply couldn't ever resolve them so we had to give up on DoubleTake. We have been using Redgate to manually read the data from the primary site (via queries) and write to the DR site but this is: A bad solution Getting the software vendor hot and bothered whenever we have support issues as it has a tendency to interfere with the payment application which is very DB intensive. We recently upgraded the whole system to run on SQL 2008 R2 Enterprise which means we should probably be looking at using some of the built-in replication features. The server has 2 fairly large databases with a mixture of tables containing highly volatile transactional data and pretty static configuration data. Replication would be done over a WAN link to a separate physical site and needs to achieve the following objectives. RPO: Zero loss - This is transactional data with financial impact so we can't lose anything. RTO: Tending to zero - The business depends on our ability to process transactions every minute we are down we are losing money I have looked at a few of the other questions/answers but none meet our case exactly: SQL Server 2008 failover strategy - Log shipping or replication? How to achieve the following RTO & RPO with logshipping only using SQL Server? What is the best of two approaches to achieve DB Replication? My current thinking is that we should use mirroring but I am concerned that for RPO:0 we will need to do delayed commits and this could impact the performance of the primary DB which is not an option. Our current DR process is to: Stop incoming traffic to the primary site and allow all in-flight transaction to complete. Allow the replication to DR to complete. Change network routing to route to DR site. Start all applications and services on the secondary site (Ideally we can change this to a warmer stand-by whereby the applications are already running but not processing any transactions). In other words the DR database needs to, as quickly as possible, catch up with primary and be ready for processing as the new primary. We would then need to be able to reverse this when we are ready to switch back. Is there a better option than mirroring (should we be doing log-shipping too) and can anyone suggest other considerations that we should keep in mind?

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  • Adding local users / passwords on Kerberized Linux box

    - by Brian
    Right now if I try to add a non-system user not in the university's Kerberos realm I am prompted for a Kerberos password anyway. Obviously there is no password to be entered, so I just press enter and see: passwd: Authentication token manipulation error passwd: password unchanged Typing passwd newuser has the same issue with the same message. I tried using pwconv in the hopes that only a shadow entry was needed, but it changed nothing. I want to be able to add a local user not in the realm and give them a local password without being bothered about Kerberos. I am on Ubuntu 10.04. Here are my /etc/pam.d/common-* files (the defaults that Ubuntu's pam-auth-update package generates): account # here are the per-package modules (the "Primary" block) account [success=1 new_authtok_reqd=done default=ignore] pam_unix.so # here's the fallback if no module succeeds account requisite pam_deny.so # prime the stack with a positive return value if there isn't one already; # this avoids us returning an error just because nothing sets a success code # since the modules above will each just jump around account required pam_permit.so # and here are more per-package modules (the "Additional" block) account required pam_krb5.so minimum_uid=1000 # end of pam-auth-update config auth # here are the per-package modules (the "Primary" block) auth [success=2 default=ignore] pam_krb5.so minimum_uid=1000 auth [success=1 default=ignore] pam_unix.so nullok_secure try_first_pass # here's the fallback if no module succeeds auth requisite pam_deny.so # prime the stack with a positive return value if there isn't one already; # this avoids us returning an error just because nothing sets a success code # since the modules above will each just jump around auth required pam_permit.so # and here are more per-package modules (the "Additional" block) # end of pam-auth-update config password # here are the per-package modules (the "Primary" block) password requisite pam_krb5.so minimum_uid=1000 password [success=1 default=ignore] pam_unix.so obscure use_authtok try_first_pass sha512 # here's the fallback if no module succeeds password requisite pam_deny.so # prime the stack with a positive return value if there isn't one already; # this avoids us returning an error just because nothing sets a success code # since the modules above will each just jump around password required pam_permit.so # and here are more per-package modules (the "Additional" block) # end of pam-auth-update config session # here are the per-package modules (the "Primary" block) session [default=1] pam_permit.so # here's the fallback if no module succeeds session requisite pam_deny.so # prime the stack with a positive return value if there isn't one already; # this avoids us returning an error just because nothing sets a success code # since the modules above will each just jump around session required pam_permit.so # and here are more per-package modules (the "Additional" block) session optional pam_krb5.so minimum_uid=1000 session required pam_unix.so # end of pam-auth-update config

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  • Simple Excel Export with EPPlus

    - by Jesse Taber
    Originally posted on: http://geekswithblogs.net/GruffCode/archive/2013/10/30/simple-excel-export-with-epplus.aspxAnyone I’ve ever met who works with an application that sits in front of a lot of data loves it when they can get that data exported to an Excel file for them to mess around with offline. As both developer and end user of a little website project that I’ve been working on, I found myself wanting to be able to get a bunch of the data that the application was collecting into an Excel file. The great thing about being both an end user and a developer on a project is that you can build the features that you really want! While putting this feature together I came across the fantastic EPPlus library. This library is certainly very well known and popular, but I was so impressed with it that I thought it was worth a quick blog post. This library is extremely powerful; it lets you create and manipulate Excel 2007/2010 spreadsheets in .NET code with a high degree of flexibility. My only gripe with the project is that they are not touting how insanely easy it is to build a basic Excel workbook from a simple data source. If I were running this project the approach I’m about to demonstrate in this post would be front and center on the landing page for the project because it shows how easy it really is to get started and serves as a good way to ease yourself in to some of the more advanced features. The website in question uses RavenDB, which means that we’re dealing with POCOs to model the data throughout all layers of the application. I love working like this so when it came time to figure out how to export some of this data to an Excel spreadsheet I wanted to find a way to take an IEnumerable<T> and just have it dumped to Excel with each item in the collection being modeled as a single row in the Excel worksheet. Consider the following class: public class Employee { public int Id { get; set; } public string Name { get; set; } public decimal HourlyRate { get; set; } public DateTime HireDate { get; set; } } Now let’s say we have a collection of these represented as an IEnumerable<Employee> and we want to be able to output it to an Excel file for offline querying/manipulation. As it turns out, this is dead simple to do with EPPlus. Have a look: public void ExportToExcel(IEnumerable<Employee> employees, FileInfo targetFile) { using (var excelFile = new ExcelPackage(targetFile)) { var worksheet = excelFile.Workbook.Worksheets.Add("Sheet1"); worksheet.Cells["A1"].LoadFromCollection(Collection: employees, PrintHeaders: true); excelFile.Save(); } } That’s it. Let’s break down what’s going on here: Create a ExcelPackage to model the workbook (Excel file). Note that the ‘targetFile’ value here is a FileInfo object representing the location on disk where I want the file to be saved. Create a worksheet within the workbook. Get a reference to the top-leftmost cell (addressed as A1) and invoke the ‘LoadFromCollection’ method, passing it our collection of Employee objects. Behind the scenes this is reflecting over the properties of the type provided and pulling out any public members to become columns in the resulting Excel output. The ‘PrintHeaders’ parameter tells EPPlus to grab the name of the property and put it in the first row. Save the Excel file All of the heavy lifting here is being done by the ‘LoadFromCollection’ method, and that’s a good thing. Now, this was really easy to do, but it has some limitations. Using this approach you get a very plain, un-styled Excel worksheet. The column widths are all set to the default. The number format for all cells is ‘General’ (which proves particularly interesting if you have a DateTime property in your data source). I’m a “no frills” guy, so I wasn’t bothered at all by trading off simplicity for style and formatting. That said, EPPlus has tons of samples that you can download that illustrate how to apply styles and formatting to cells and a ton of other advanced features that are way beyond the scope of this post.

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  • Simple Scripting for your Exalogic Storage

    - by Trond Strømme
    As part of my job in Oracle ACS (Advanced Customer Services) I'm handling lots of different systems and customers. Among the recent systems I worked with have been Oracle's Exalogic engineered systems. One of the things I'd never had much exposure to as a system developer/architect/middleware guy/Java dude has been storage; outside of consuming it for my photography needs.. Well, I'm always ready for a new challenge... I'd downloaded the 7000 series storage simulator when it was released in the good old Sun days, found it fun and instructive to play around with, but as I never touched storage in any way (besides consuming it..) I forgot about it. A couple of years ago when I started working with Exalogic engineered systems it again came into light as an invaluable learning and testing tool for the embedded storage in an Exalogic;  Oracle's Sun ZFS Storage 7320 Appliance.  aaaanyway... I've been "booted" into a part-time role as the interim storage/system admin/middleware/Java guy for a client and found I needed to create the occasional report or summary or whatever.. of what's using the storage in the 7320 (as default configured for an Exalogic, 40T of disk in a mirrored configuration, yielding 18T of actual space.) Reading the nice documentation and some articles on the Oracle Technology Network I saw great possibilities with the embedded ECMAScript3/JavaScript engine in the 7000 series.  In my personal opinion anyone who's dealing with Exalogic administration, or exposed to any of the 7000 series of storage appliances and servers that Oracle offers should have a VirtualBox instance of it kicking around. For development and testing it's a fantastic tool. (It can save you from explaining (most) of the embarrassing FAILS you can do if you test something in a production system to your management...) So download, and install.  A small sidestep, if after firing up the 7000 series simulator in VirtualBox you've forgotten what it's IP address is, the following will sort you out if you log in directly via the running VirtualBox VM. So in my case I can ssh to 192.168.56.101 or point a browser to https://192.168.56.101:215 to log into the storage appliance. One simple way of executing a script on the 7320 is to ssh to the device and redirecting a file with the script in it to ssh. ssh [email protected] < myscript.js One question I got from my client and the people who will take over the systems was: "how can we see the quotas and allocations for all projects/shares in one easy go so we don't have to go navigating around in the BUI for all the hundreds of shares the 7320 is hosting just to check if anything is running dry?" Easy! JavaScript time, VirtualBox and emacs! //NOTE! this script is available 'as is' It has ben run on a couple of 7320's, (running 2010.08.17.3.0,1-1.25 & // 2011.04.24.1.0,1-1.8) a 7420 and the VB image, but I personally //offer no guarantee whatsoever that it won't make your server topple, catch fire or in any way go pear shaped.. //run at your own risk or learn from my code and or mistakes.. script run('cd /'); run('shares'); //get all projects: proj = list(); function spaceToGig(bytes){ return bytes/1073741824; //convert bytes to GB } function fullInPercent(quota, space_data){ tmp = (space_data/quota)*100; return tmp; } //print header, slightly good looking printf(" %s/%-15s %8s(GB) %7s(GB) %5s(GB) %7s(GB) %3s\n","Project", "Share","Quota","Ref", "Snap", "Total","%full"); printf("-------------------------------------------------------------------------------\n") //for each project, get all shares. check for quota and calculate percentage and human readable figures.. for (i=0;i<proj.length;i++){ run('select ' + proj[i]); //get all shares for a project var pshares = list(); //for each share get quota properties for (j=0;j<pshares.length;j++){ run('select ' + pshares[j]); quota = get('quota'); //properties associated with a share or inherited from a project spaceData = get('space_data'); spaceSnap = get('space_snapshots'); spaceTotal = get('space_total'); if(quota>0){ //has quota printf(" %s/%-15s \t%4.2fGB\t%.2fGB\t%.2fGB\t%.2fGB\t%5.2f%%\n",proj[i], pshares[j],spaceToGig(quota),spaceToGig(spaceData),spaceToGig(spaceSnap),spaceToGig(spaceTotal),fullInPercent(quota,spaceTotal)); }else{ //no quota printf(" %s/%-15s \t%8s\t%.2fGB\t%.2fGB\t%.2fGB\t%s\n",proj[i],pshares[j], "N/A", spaceToGig(spaceData),spaceToGig(spaceSnap),spaceToGig(spaceTotal),"N/A"); } run('cd ..'); } run('done'); } The resulting output should look something like this: Project/Share Quota(GB) Ref(GB) Snap(GB) Total(GB) %full ------------------------------------------------------------------------------- ACSExalogicSystem/domains N/A 0.04GB 0.00GB 0.04GB N/A ACSExalogicSystem/logs N/A 0.01GB 0.00GB 0.01GB N/A ACSExalogicSystem/nodemgrs N/A 0.00GB 0.00GB 0.00GB N/A ACSExalogicSystem/stores N/A 0.04GB 0.00GB 0.04GB N/A ***_dev/FMW_***_1 133GB 4.24GB 0.01GB 4.25GB 3.19% ***_dev/FMW_***_2 N/A 4.25GB 0.01GB 4.26GB N/A ***_dev/applications 10GB 0.00GB 0.00GB 0.00GB 0.00% ***_dev/domains 50GB 10.75GB 3.55GB 14.30GB 28.61% ***_dev/logs 20GB 0.32GB 0.01GB 0.33GB 1.66% ***_dev/softwaredepot 20GB 4.15GB 0.00GB 4.15GB 20.73% ***_dev/stores 20GB 0.01GB 0.00GB 0.01GB 0.05% ###_dev/FMW_###_1 400GB 17.63GB 0.12GB 17.75GB 4.44% ###_dev/applications N/A 0.00GB 0.00GB 0.00GB N/A ###_dev/domains 120GB 14.21GB 5.53GB 19.74GB 16.45% ###_dev/logs 15GB 0.00GB 0.00GB 0.00GB 0.00% ###_dev/softwaredepot 250GB 73.55GB 0.02GB 73.57GB 29.43% …snip My apologies if the output is a bit mis-aligned here and there, I only bothered making it look good, not perfect :/ I also removed some of the project names (*,#)

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  • Reading the tea leaves from Windows Azure support

    - by jamiet
    A few idle thoughts… Three months ago I had an issue regarding Windows Azure where I was unable to login to the management portal. At the time I contacted Azure support, the issue was soon resolved and I thought no more about it. Until today that is when I received an email from Azure support providing a detailed analysis of the root cause, the fix and moreover precise details about when and where things occurred. The email itself is interesting and I have included the entirety of it below. A few things were interesting to me: The level of detail and the diligence in investigating and reporting the issue I found really rather impressive. They even outline the number of users that were affected (127 in case you can’t be bothered reading). Compare this to the quite pathetic support that another division within Microsoft, Skype, provided to Greg Low recently: Skype support and dead parrot sketches   This line: “Windows Azure performed a planned change from using the Microsoft account service (formerly Windows Live ID) to the Azure Active Directory (AAD) as its primary authentication mechanism on August 24th. This change was made to enable future innovation in the area of authentication – particularly for organizationally owned identities, identity federation, stronger authentication methods and compliance certification. ” I also found to be particularly interesting. I have long thought that one of the reasons Microsoft has proved to be such a money-making machine in the enterprise is because they provide the infrastructure and then upsell on top of that – and nothing is more infrastructural than Active Directory. It has struck me of late that they are trying to make the same play of late in the cloud by tying all their services into Azure Active Directory and here we see a clear indication of that by making AAD the authentication mechanism for anyone using Windows Azure. I get the feeling that we’re going to hear much much more about AAD in the future; isn’t it about time we could log on to SQL Azure Windows Azure SQL Database without resorting to SQL authentication, for example? And why do Microsoft have two identity providers – Microsoft Account (aka Windows Live ID) and AAD – isn’t it about time those things were combined? As I said, just some idle thoughts. Below is the transcript of the email if you are interested. @Jamiet  This is regarding the support request <redacted> where in you were not able to login into the windows azure management portal with live id. We are providing you with the summary, root cause analysis and information about permanent fix: Incident Title: You were unable to access Windows Azure Portal after Microsoft Account to Azure Active Directory account Migration. Service Impacted: Management Portal Incident Start Date and Time: 8/24/2012 4:30:00 PM Date and Time Service was Restored: 10/17/2012 12:00:00 AM Summary: Windows Azure performed a planned change from using the Microsoft account service (formerly Windows Live ID) to the Azure Active Directory (AAD) as its primary authentication mechanism on August 24th.   This change was made to enable future innovation in the area of authentication – particularly for organizationally owned identities, identity federation, stronger authentication methods and compliance certification.   While this migration was largely transparent to Windows Azure users, a small number of users whose sign-in names were part of a Windows Live Custom Domain were unable to login.   This incompatibility was not discovered during the Quality Assurance testing phase prior to the migration. Customer Impact: Customers whose sign-in names were part of a Windows Live Custom Domain were unable to sign-in the Management Portal after ~4:00 p.m. PST on August 24th, 2012.   We determined that the issue did impact at least 127 users in 98 of these Windows Live Custom Domains and had a maximum potential impact of 1,110 users in total. Root Cause: The root cause of the issue was an incompatibility in the AAD authentication service to handle logins from Microsoft accounts whose sign-in names were part of a Windows Live Custom Domains.  This issue was not discovered during the Quality Assurance testing phase prior to the migration from Microsoft Account (MSA) to AAD. Mitigations: The issue was mitigated for the majority of affected users by 8:20 a.m. PST on August 25th, 2012 by running some internal scripts to correct many known Windows Live Custom Domains.   The remaining affected domains fell into two categories: Windows Live Custom Domains that were not corrected by 8/25/2012. An additional 48 Windows Live Custom Domains were fixed in the weeks following the incident within 2 business days after the AAD team received an escalation from product support regarding those accounts. Windows Live Custom domains that were also provisioned in Office365. Some of the affected Windows Live Custom Domains had already been provisioned in AAD because their owners signed up for Office365 which is a service that also uses AAD.   In these cases the Azure customers had to work around the issue by renaming their Microsoft Account or using a different Microsoft Account to administer their Azure subscription. Permanent Fix: The Azure Active Directory team permanently fixed the issue for all customers on 10/17/2012 in an upgraded release of the AAD service.

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  • Part 4 of 4 : Tips/Tricks for Silverlight Developers.

    - by mbcrump
    Part 1 | Part 2 | Part 3 | Part 4 I wanted to create a series of blog post that gets right to the point and is aimed specifically at Silverlight Developers. The most important things I want this series to answer is : What is it?  Why do I care? How do I do it? I hope that you enjoy this series. Let’s get started: Tip/Trick #16) What is it? Find out version information about Silverlight and which WebKit it is using by going to http://issilverlightinstalled.com/scriptverify/. Why do I care? I’ve had those users that its just easier to give them a site and say copy/paste the line that says User Agent in order to troubleshoot a Silverlight problem. I’ve also been debugging my own Silverlight applications and needed an easy way to determine if the plugin is disabled or not. How do I do it: Simply navigate to http://issilverlightinstalled.com/scriptverify/ and hit the Verify button. An example screenshot is located below: Results from Chrome 7 Results from Internet Explorer 8 (With Silverlight Disabled) Tip/Trick #17) What is it? Use Lambdas whenever you can. Why do I care?  It is my personal opinion that code is easier to read using Lambdas after you get past the syntax. How do I do it: For example: You may write code like the following: void MainPage_Loaded(object sender, RoutedEventArgs e) { //Check and see if we have a newer .XAP file on the server Application.Current.CheckAndDownloadUpdateAsync(); Application.Current.CheckAndDownloadUpdateCompleted += new CheckAndDownloadUpdateCompletedEventHandler(Current_CheckAndDownloadUpdateCompleted); } void Current_CheckAndDownloadUpdateCompleted(object sender, CheckAndDownloadUpdateCompletedEventArgs e) { if (e.UpdateAvailable) { MessageBox.Show( "An update has been installed. To see the updates please exit and restart the application"); } } To me this style forces me to look for the other Method to see what the code is actually doing. The style located below is much easier to read in my opinion and does the exact same thing. void MainPage_Loaded(object sender, RoutedEventArgs e) { //Check and see if we have a newer .XAP file on the server Application.Current.CheckAndDownloadUpdateAsync(); Application.Current.CheckAndDownloadUpdateCompleted += (s, e) => { if (e.UpdateAvailable) { MessageBox.Show( "An update has been installed. To see the updates please exit and restart the application"); } }; } Tip/Trick #18) What is it? Prevent development Web Service references from breaking when Visual Studio auto generates a new port number. Why do I care?  We have all been there, we are developing a Silverlight Application and all of a sudden our development web services break. We check and find out that the local port number that Visual Studio assigned has changed and now we need up to update all of our service references. We need a way to stop this. How do I do it: This can actually be prevented with just a few mouse click. Right click on your web solution and goto properties. Click the tab that says, Web. You just need to click the radio button and specify a port number. Now you won’t be bothered with that anymore. Tip/Trick #19) What is it? You can disable the Close Button a ChildWindow. Why do I care?  I wouldn’t blog about it if I hadn’t seen it. Devs trying to override keystrokes to prevent users from closing a Child Window. How do I do it: A property exist on the ChildWindow called “HasCloseButton”, you simply change that to false and your close button is gone. You can delete the “Cancel” button and add some logic to the OK button if you want the user to respond before proceeding. Tip/Trick #20) What is it? Cleanup your XAML. Why do I care?  By removing unneeded namespaces, not naming all of your controls and getting rid of designer markup you can improve code quality and readability. How do I do it: (This is a 3 in one tip) Remove unused Designer markup: 1) Have you ever wondered what the following code snippet does? xmlns:d="http://schemas.microsoft.com/expression/blend/2008" xmlns:mc="http://schemas.openxmlformats.org/markup-compatibility/2006" mc:Ignorable="d" d:DesignWidth="640" d:DesignHeight="480" This code is telling the designer to do something special with this page in “Design mode” Specifically the width and the height of the page. When its running in the browser it will not use this information and it is actually ignored by the XAML parser. In other words, if you don’t need it then delete it. 2) If you are not using a namespace then remove it. In the code sample below, I am using Resharper which will tell me the ones that I’m not using by the grayed out line below. If you don’t have resharper you can look in your XAML and manually remove the unneeded namespaces. 3) Don’t name an control unless you actually need to refer to it in procedural code. If you name a control you will take a slight performance hit that is totally unnecessary if its not being called. <TextBlock Height="23" Text="TextBlock" />   That is the end of the series. I hope that you enjoyed it and please check out Part 1 | Part 2 | Part 3 if your hungry for more.  Subscribe to my feed CodeProject

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  • Opinions on sensor / reading / alert database design

    - by Mark
    I've asked a few questions lately regarding database design, probably too many ;-) However I beleive I'm slowly getting to the heart of the matter with my design and am slowly boiling it down. I'm still wrestling with a couple of decisions regarding how "alerts" are stored in the database. In this system, an alert is an entity that must be acknowledged, acted upon, etc. Initially I related readings to alerts like this (very cut down) : - [Location] LocationId [Sensor] SensorId LocationId UpperLimitValue LowerLimitValue [SensorReading] SensorReadingId Value Status Timestamp [SensorAlert] SensorAlertId [SensorAlertReading] SensorAlertId SensorReadingId The last table is associating readings with the alert, because it is the reading that dictate that the sensor is in alert or not. The problem with this design is that it allows readings from many sensors to be associated with a single alert - whereas each alert is for a single sensor only and should only have readings for that sensor associated with it (should I be bothered that the DB allows this though?). I thought to simplify things, why even bother with the SensorAlertReading table? Instead I could do this: [Location] LocationId [Sensor] SensorId LocationId [SensorReading] SensorReadingId SensorId Value Status Timestamp [SensorAlert] SensorAlertId SensorId Timestamp [SensorAlertEnd] SensorAlertId Timestamp Basically I'm not associating readings with the alert now - instead I just know that an alert was active between a start and end time for a particular sensor, and if I want to look up the readings for that alert I can do. Obviously the downside is I no longer have any constraint stopping me deleting readings that occurred during the alert, but I'm not sure that the constraint is neccessary. Now looking in from the outside as a developer / DBA, would that make you want to be sick or does it seem reasonable? Is there perhaps another way of doing this that I may be missing? Thanks. EDIT: Here's another idea - it works in a different way. It stores each sensor state change, going from normal to alert in a table, and then readings are simply associated with a particular state. This seems to solve all the problems - what d'ya think? (the only thing I'm not sure about is calling the table "SensorState", I can't help think there's a better name (maybe SensorReadingGroup?) : - [Location] LocationId [Sensor] SensorId LocationId [SensorState] SensorStateId SensorId Timestamp Status IsInAlert [SensorReading] SensorReadingId SensorStateId Value Timestamp There must be an elegant solution to this!

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  • Using OpenID as the only authentication method

    - by iconiK
    I have read the other questions and they mostly talk about the security of doing so. That's not entirely my concern, mostly because the website is question is a browser-based game. However, the larger issue is the user - not every user is literate enough to understand OpenID. Sure RPX makes this pretty easy, which is what I'll use, but what if the user does not have an account at Google or Facebook or whatever, or does not trust the system to log in with an existing account? They'd have to get an account at another provide - I'm sure most will know how to do it, let alone be bothered to do it. There is also the problem of how to manage it in the application. A user might want to use multiple identities with a single account, so it's not as simple as username + password to deal with. How do I store the OpenID identities of a user in the database? Using OpenID gives me a benefit too: RPX can provide extensive profile information, so I can just prefill the profile form and ask the user to edit as required. I currently have this: UserID Email ------ --------------- 86000 [email protected] 86001 [email protected] UserOpenID OpenID ---------- ------ 86000 16733 86001 16839 86002 19361 OpenID Provider Identifier ------ -------- ---------------- 16733 Yahoo https:\\me.yahoo.com\bob#d36bd 16839 Yahoo https:\\me.yahoo.com\bigbobby#x75af 19361 Yahoo https:\\me.yahoo.com\alice#c19fd Is that the right way to store OpenID identifiers in the database? How would I match the identifier RPX gave me with one in the database to log in the user (if the identifier is known). So here are concrete questions: How would I make it accessible to users not having an OpenID or not wanting to use one? (security concerns over say, logging in with their Google account for example) How do I store the identifier in the database? (I'm not sure if the tables above are right) What measures do I need to take in order to prevent someone from logging in as another user and happily doing anything with their account? (as I understand RPX sends the identifier via HTTP, so what anyone would have to do is to just somehow grab it then enter it in the "OpenID" field) What else do I need to be aware of when using OpenID?

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  • Referencing both an old version and new version of the same DLL (VB.Net)

    - by ckittel
    Consider the following situation: WidgetCompany produced a .NET DLL in 2006 called Widget.dll, version 1.0. I consumed this Widget.dll file throughout my VB.Net application. Over time, WidgetCompany has been updating Widget.dll, I never bothered to keep up, continuing to ship version 1.0 of Widget.dll with my software. It's now 2010, my project is now a VB.Net 3.5 application and WidgetCompany has come out with Widget.dll version 3.0. It looks and functions almost identical to Widget.dll version 1.0, using all the same namespaces and type names from before. However, Widget.dll version 3.0 has many run-time breaking changes since 1.0 and I cannot simply cut over to the new version; however, I don't want to continue developing against the 1.0 version and therefore keep digging myself deeper in the hole. What I want to do is do all new development in my project with Widget.dll version 3.0, whilst keeping Widget.dll version 1.0 around until I find time to convert all of my 1.0 consumption to the newer 3.0 code. Now, for starters, I obviously cannot simply reference both Widget.dll (Ver 1.0) and Widget.dll (Ver 3.0) in Visual Studio. Doing so gives me the following message: "A reference to 'Widget.dll' could not be added. A reference to the component 'Widget' already exists in the project." To work around that, I can simply rename version 3.0 Widget.dll to Widget.3.dll. But this is where I'm stuck. Any attempts to reference types found in "the dll" leads to ambiguity and the compiler obviously doesn't have any clue as to what I really want in this or that case. Is there something I can do that gives a DLL a new "root" Namespace or something? For example, if I could say "Widget.dll has a new root namespace of Legacy" then I could update existing code to reference the types found in Legacy.<RootNamespace> namespace while all new code could simply reference types from the <RootNamespace> namespace. Pipe dream or reality? Are there other solutions to situations this (besides "don't get in this situation in the first place")?

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  • Function-Local Static Const variable Initialization semantics.

    - by Hassan Syed
    The questions are in bold, for those that cannot be bothered reading a question in depth. This is a followup to this question. It is to do with the initialization semantics of static variables in functions. Static variables should be initialized once, and their internal state might be altered later - as I (currently) do in the linked question. However, the code in question does not require the feature to change the state of the variable later. Let me clarrify my position, since I don't require the string object's internal state to change. The code is for a trait class for meta programming, and as such would would benifit from a const char * const ptr -- thus Ideally a local cost static const variable is needed. My educated guess is that in this case the string in question will be optimally placed in memory by the link-loader, and that the code is more secure and maps to the intended semantics. This leads to the semantics of such a variable "The C++ Programming language Third Edition -- Stroustrup" does not have anything (that I could find) to say about this matter. All that is said is that the variable is initialized once when the flow of control of the thread first reaches the code. This leads me to ponder if the following code would be sensible, and if not what are the intended semantics ?. #include <iostream> const char * const GetString(const char * x_in) { static const char * const x = x_in; return x; } int main() { const char * const temp = GetString("yahoo"); std::cout << temp << std::endl; const char * const temp2 = GetString("yahoo2"); std::cout << temp2 << std::endl; } The following compiles on GCC and prints "yahoo" twice. Which is what I want -- However it might not be standards compliant (which is why I post this question). It might be more elegant to have two functions, "SetString" and "String" where the latter forwards to the first. If it is standards compliant does someone know of a templates implementation in boost (or elsewhere) ?

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  • Can you help me optimize this code for finding factors of a number? I'm brushing up on my math progr

    - by Sergio Tapia
    I've never really bothered with math programming, but today I've decided to give it a shot. Here's my code and it's working as intended: using System; using System.Collections.Generic; using System.Linq; using System.Text; using System.Windows; using System.Windows.Controls; using System.Windows.Data; using System.Windows.Documents; using System.Windows.Input; using System.Windows.Media; using System.Windows.Media.Imaging; using System.Windows.Navigation; using System.Windows.Shapes; namespace PrimeFactorization { /// <summary> /// Interaction logic for MainWindow.xaml /// </summary> public partial class MainWindow : Window { public MainWindow() { InitializeComponent(); } private void btnSubmit_Click(object sender, RoutedEventArgs e) { List<int> primeFactors = FindPrimeFactors(Convert.ToInt32(txtNumber.Text)); primeFactors.Sort(); for (int i = 0; i < primeFactors.Count; i++) { listBoxFoundNumbers.Items.Add(primeFactors[i]); } } private List<int> FindPrimeFactors(int number) { List<int> factors = new List<int>(); factors.Add(1); factors.Add(number); for (int i = 2; i < number; i++) { if (number % i == 0) { int holder = number / i; //If the number is in the list, don't add it again. if (!factors.Contains(i)) { factors.Add(i); } //If the number is in the list, don't add it again. if (!factors.Contains(holder)) { factors.Add(holder); } } } return factors; } } } The only problem I can see with my code is that it will iterate through to the bitter end, even though there will definitely not be any factors. For example, imagine I wrote in 35. My loop will go up to 35 and check 24,25,26,27...etc. Not very good. What do you recommend?

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  • JSON or YAML encoding in GWT/Java on both client and server

    - by KennethJ
    I'm looking for a super simple JSON or YAML library (not particularly bothered which one) written in Java, and can be used in both GWT on the client, and in its original Java form on the server. What I'm trying to do is this: I have my models, which are shared between the client and the server, and these are the primary source of data interchange. I want to design the web service in between to be as simple as possible, and decided to take the RESTful approach. My problem is that I know our application will grow substantially in the future, and writing all the getters, setters, serialization, factories, etc. by hand fills me with absolute dread. So in order to avoid it, I decided to implement annotations to keep track of attributes on the models. The reason I can't just serialize everything directly, using GWT's own one, or one which works through reflection, is because we need a certain amount of logic going on in the serialization process. I.e. whether references to other models get serialized during the serialization of the original model, or whether an ID is just passed, and general simple things like that. I've then written an annotation processor to preprocess my shared models and generate an implementing class with all the getters, setters, serialization, lazy-loading, etc. To make a long story short, I need some type of simple YAML or JSON library, which allows me to encode and decode manually, so I can generate this code through my annotation processor. I have had a look around the interwebs, but every single one I ran into supported some reflection which, while all fine and dandy, make it pretty much useless for GWT. And in the case of GWT's own JSON library, it uses JSNI for speed purposes, making it useless server side. One solution I did think about involved writing writing two sets of serialization methods on the models, one for the client and one for the server, but I'd rather not do that. Also, I'm pretty new to GWT, and even though I have done a lot of Java, it was back in the 1.2 days, so it's a bit rusty. So if you think I'm going about this problem completely the wrong way, I'm open to suggestions.

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  • Design patterns and interview question

    - by user160758
    When I was learning to code, I read up on the design patterns like a good boy. Long after this, I started to actually understand them. Design discussions such as those on this site constantly try to make the rules more and more general, which is good. But there is a line, over which it becomes over-analysis starts to feed off itself and as such I think begins to obfuscate the original point - for example the "What's Alternative to Singleton" post and the links contained therein. http://stackoverflow.com/questions/1300655/whats-alternative-to-singleton I say this having been asked in both interviews I’ve had over the last 2 weeks what a singleton is and what criticisms I have of it. I have used it a few times for items such as user data (simple key-value eg. last file opened by this user) and logging (very common i'm sure). I've never ever used it just to have what is essentially global application data, as this is clearly stupid. In the first interview, I reply that I have no criticisms of it. He seemed disappointed by this but as the job wasn’t really for me, I forgot about it. In the next one, I was asked again and, as I wanted this job, I thought about it on the spot and made some objections, similar to those contained in the post linked to above (I suggested use of a factory or dependency injection instead). He seemed happy with this. But my problem is that I have used the singleton without ever using it in this kind of stupid way, which I had to describe on the spot. Using it for global data and the like isn’t something I did then realised was stupid, or read was stupid so didn’t do, it was just something I knew was stupid from the start. Essentially I’m supposed to be able to think of ways of how to misuse a pattern in the interview? Which class of programmers can best answer this question? The best ones? The medium ones? I'm not sure.... And these were both bright guys. I read more than enough to get better at my job but had never actually bothered to seek out criticisms of the most simple of the design patterns like this one. Do people think such questions are valid and that I ought to know the objections off by heart? Or that it is reasonable to be able to work out what other people who are missing the point would do on the fly? Or do you think I’m at least partially right that the question is too unsubtle and that the questions ought to be better thought out in order to make sure only good candidates can answer. PS. Please don’t think I’m saying that I’m just so clever that I know everything automatically - I’ve learnt the hard way like everyone else. But avoiding global data is hardly revolutionary.

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  • Security review of an authenticated Diffie Hellman variant

    - by mtraut
    EDIT I'm still hoping for some advice on this, i tried to clarify my intentions... When i came upon device pairing in my mobile communication framework i studied a lot of papers on this topic and and also got some input from previous questions here. But, i didn't find a ready to implement protocol solution - so i invented a derivate and as i'm no crypto geek i'm not sure about the security caveats of the final solution: The main questions are Is SHA256 sufficient as a commit function? Is the addition of the shared secret as an authentication info in the commit string safe? What is the overall security of the 1024 bit group DH I assume at most 2^-24 bit probability of succesful MITM attack (because of 24 bit challenge). Is this plausible? What may be the most promising attack (besides ripping the device out off my numb, cold hands) This is the algorithm sketch For first time pairing, a solution proposed in "Key agreement in peer-to-peer wireless networks" (DH-SC) is implemented. I based it on a commitment derived from: A fix "UUID" for the communicating entity/role (128 bit, sent at protocol start, before commitment) The public DH key (192 bit private key, based on the 1024 bit Oakley group) A 24 bit random challenge Commit is computed using SHA256 c = sha256( UUID || DH pub || Chall) Both parties exchange this commitment, open and transfer the plain content of the above values. The 24 bit random is displayed to the user for manual authentication DH session key (128 bytes, see above) is computed When the user opts for persistent pairing, the session key is stored with the remote UUID as a shared secret Next time devices connect, commit is computed by additionally hashing the previous DH session key before the random challenge. For sure it is not transfered when opening. c = sha256( UUID || DH pub || DH sess || Chall) Now the user is not bothered authenticating when the local party can derive the same commitment using his own, stored previous DH session key. After succesful connection the new DH session key becomes the new shared secret. As this does not exactly fit the protocols i found so far (and as such their security proofs), i'd be very interested to get an opinion from some more crypto enabled guys here. BTW. i did read about the "EKE" protocol, but i'm not sure what the extra security level is.

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  • creating a heirarchy of terminals or workspaces

    - by intuited
    <rant This question occurred to me ('occurred' meaning 'whispered seductively in my ear for the 100th time') while using GNU-screen, so I'll make that my example. However this is a much more general question about user interfaces and what I perceive as a flawmissing feature in every implementation I've yet seen. I'm wondering if there is some way to create a heirarchy/tree of terminals in a screen session. EG I'd like to have something like 1 bash 1.1 bash 1.2 bash 2 bash 3 bash 3.1 bash 3.1.1 bash 3.1.2 bash It would be good if the terminals could be labelled instead of having to be navigated to via some arrangement that I suspect doesn't exist. So then you could jump to one using eg ^A:goto happydays or ^A:goto dykstra.angry. So to generalize that: Firefox, Chrome, Internet Explorer, gnome-terminal, roxterm, konsole, yakuake, OpenOffice, Microsoft Office, Mr. Snuffaluppagus's Funtime Carousel™, and Your Mom's Jam Browser™ all offer the ability to create a flat set of tabs containing documents of an identical nature: web pages, terminals, documents, fun rideable animals, and jams. GNU-screen implements the same functionality without using tabs. Linux and OS/X window managers provide the ability to organize windows into an array of workspaces, which amounts to again, the same deal. Over the past few years, this has become a more or less ubiquitous concept which has been righteously welcomed into the far reaches of the computer interface funfest. Heavy users of these systems quickly encounter a problem with it: the set of entities is flat. In the case of workspaces, an option may be available to create a 2d array. However none of these applications furnish their users with the ability to create heirarchies, similar to filesystem directory structures, containing instances of their particular contained type. I for one am consistently bothered by this, and am wondering if the community can offer some wisdom as to why this has not happened in any of the foremost collections of computational functionality our culture has yet produced. Or if perhaps it has and I'm just an ignorant savage. I'd like to be able to not only group things into a tree structure, but also to create references (aka symbolic links, aka pointers) from one part of the structure to another, as well as apply properties (eg default directory, colorscheme, ...) recursively downward from a given node. I see no reason why we shouldn't be able to save these structures as known sessions, and apply tags to particular instances. So then you can sort through them by tag, find them by name, or just use the arrow keys (with an appropriate modifier) to move left or right and in or out of a given level. Another key combo would serve to create a branch in the place of the current terminal/webpage/lifelike statue/spreadsheet/spreadsheet sheet/presentation/jam and move that entity into the new branch, then create a fresh one as a sibling to it: a second leaf node within the same branch node. They would get along well. I find it a bit astonishing that this hasn't happened yet, and the only reason I can venture as a guess is that the creators of these fine systems do not consider such functionality to be useful to a significant portion of their userbase. I posit that the probability that that such an assumption would be correct is pretty low. On the other hand, given the relative ease with which such structures can be implemented using modern libraries/languages, it doesn't seem likely that difficulty of implementation would be a major roadblock. If it could be done in 1972 or whenever within the constraints of a filesystem driver, it should be relatively painless to implement in 2010 in a fullblown application. Given that all of these systems are capable of maintaining a set of equivalent entities, it seems unlikely that a major infrastructure overhaul would be necessary in order to enable a navigable heirarchy of them. </rant Mostly I'm just looking to start up a discussion and/or brainstorming on this topic. Any ideas, examples, criticism, or analysis are quite welcome. * Mr. Snuffaluppagus's Funtime Carousel is a registered trademark of Children's Television Workshop Inc. * Your Mom's Jam Browser is a registered trademark of Your Mom Inc.

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  • Understanding NFS4 (Linux server)

    - by drumfire
    I've been a bit bothered by NFS4 on Linux. Some information 'out there' seems to conflict with other information, and other information appears hard to find. So here are a couple of things that caught my attention, hopefully someone out there can shed some light on this. This question focuses exclusively on NFS4 without Kerberos etc. 1. Exports There is ambiguous information in the exports manpage on the structure of /etc/exports. To quote from exports(5): Also, each line may have one or more specifications for default options after the path name, in the form of a dash ("-") followed by an option list. The option list is used for all subsequent exports on that line only. What does "subsequent exports on that line only" mean? 1.2 fsid=0 not required anymore? I was searching for fsid when I found a comment on the linux-nfs list stating fsid=0 is not required anymore. Now I'm just confused, do I need it with nfs4 or not?! 2. Non-exported directory still mountable Say I have the following tree: /exp /exp/users /exp/distr /exp/distr/archlinux /exp/distr/debian And I have the following entries in this fstab entry: /dev/disk/by-label/users /mnt/users ext4 defaults 0 0 /dev/disk/by-label/distr /mnt/distr ext4 defaults 0 0 /mnt/users /exp/users none bind 0 0 /mnt/distr /exp/distr none bind 0 0 And my exports is exactly this: /exp 192.168.1.0/24(fsid=0,rw,async,no_subtree_check,no_root_squash) /exp/distr 192.168.1.0/24(rw,async,no_subtree_check,no_root_squash) And exportfs -arv shows: exporting 192.168.1.0/24:/exp/distr exporting 192.168.1.0/24:/exp Then why am I able to do this and get no error on a client: mount -t nfs4 server:/exp/users /tmp/test Even though /exp/users is not exported? I didn't export this directory, and while I don't see the contents of /dev/disk/by-label/users unless I specify crossmnt, I am still able to write to the directory. Everything I write to there goes to the underlying directory of /exp/users which can be seen when I umount /exp/users; ls /exp/users.. 3. The odd case of showmount -d server As stated by rpc.mountd(8), this command should display directories that are either currently mounted by clients, or stale entries in /var/lib/nfs/rmtab, as can be read: The rpc.mountd daemon registers every successful MNT request by adding an entry to the /var/lib/nfs/rmtab file. When receivng a UMNT request from an NFS client, rpc.mountd simply removes the matching entry from /var/lib/nfs/rmtab, as long as the access control list for that export allows that sender to access the export. (...) Note, however, that there is little to guarantee that the contents of /var/lib/nfs/rmtab are accurate. A client may continue accessing an export even after invoking UMNT. If the client reboots without sending a UMNT request, stale entries remain for that client in /var/lib/nfs/rmtab. After reading this I surely wonder: Isn't it terribly insecure to just expose this type of client information; Aren't unaware server admins bound to have an rmtab with a lot of stale clients; Is this the reason that clients that mount nfs4 directories with mount -v get to see output like "nothing was mounted" even though something was mounted? I have a lot of other questions regarding nfs4, but I'll keep it at this for the moment.. :)

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  • RAID controller dropping the wrong drive

    - by bramp
    I've been having an issue with 3ware 9500S-8 RAID 10, and I have contracted their tech support, but I wanted to hear the serverfault community's recommendations. Firstly, all my data is backuped and secure, so I don't mind blowing my RAID away if I have to. But let me describe the problem I've been seeing. A month ago, disk 6 dropped out of the RAID. It is mirrored with disk 7, so I wasn't that bothered. I went to the data centre and replaced it. When I got back to the office, I noticed that disk 6 will still not in the RAID, and in fact the controller was show the name of the old drive still. A week later I went back and replace the drive again, thinking I might have swapped in a bad drive. Still the same problem. I decided to reboot the machine, to see if that would "force" the controller into seeing the new drive. It did, and a rebuild started to happen (from disk 7). Eventually both drives were showing as good. A week later, the MySQL database has flagged the database is corrupt, and is unable to repair it. I don't know what has gone wrong, but I suspected this 6-7 pair. At this point I noticed that the RAID had constantly been verifying itself, over and over. Regardless of this I began to rebuild the database, which took about 19 hours. It's a big database. Near the end of the repair, the RAID controller told me it had dropped disk 7, and that some data was most likely corrupted. I contacted LSI tech support, and they very promptly started to help me. I mentioned that drive 7 had been dropped. They suspect that drive 7 was always at fault, and drive 6 had always been good. I want to know how often a RAID controller would drop the wrong drive (in this case dropping drive 6 a month ago, instead of 7). I foolishly didn't run smartctl on the drives before I started swapping them out. I just assumed the RAID controller knew what it was talking about. I think my plan of action is to replace drive 7, rebuild the array from scratch, double check smartctl on ALL the disks, and then start restoring my data again. I would appreciate anyone's input on what the correct procedure for swapping drives is, and how often failures like this happen. If anyone would like more information then I'd be happy to provide it. thanks in advance. Oh some more information. I'm running CentOS 5.3, with two RAID arrays, a simple RAID 1 for the OS, and RAID 10 for the database. Both arrays are on different controllers. The RAID 10 is made of 10 identical ST3640323AS drives, until I swapped in a SAMSUNG HD103SJ last month.

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  • Specifying a Postfix Instance to send outbound email

    - by Catherine Jefferson
    I have a CentOS 6.5 server running Postfix 2.6x (the default distribution) with five public IPv4 IPs bound to it. Each IP has DNS and rDNS set separately. Each uses a different hostname at a different domain. I have five Postfix instances, one bound to each IP, like this example: 192.168.34.104 red.example.com /etc/postfix 192.168.36.48 green.example.net /etc/postfix-green 192.168.36.49 pink.example.org /etc/postfix-pink 192.168.36.50 orange.example.info /etc/postfix-orange 192.168.36.51 blue.example.us /etc/postfix-blue I've tested each IP by telneting to port 25. Postfix answers and banners properly with the correct hostname. Email is received on all of these instances with no problems and is routed to the correct place. This setup, minus the final instance, has existed for a couple of years and works. I never bothered to set up outbound email to go through any but the main instance, however; there was no need. Now I need to send email from blue.example.us that actually leaves from that interface and IP, such that the Received headers show blue.example.us as the sending mailhost, so that SPF and DKIM validate, etc etc. The email that will be sent from blue.example.com is a feedback loop sent by a single shell account on the server (account5), an account that is dedicated to sending this email. The account receives the feedback loop emails from servers on other networks, saves the bodies of those emails, and then generates a new outbound email header, appends the saved body, and sends the email. It's sending by piping each email to sendmail -oi -t. We're doing it this way to mask the identities of the initial servers. The procmail script that processes these emails works correctly. However, I cannot configure this account to send email through the proper Postfix instance/IP/interface. The exact same account and script sends email through the main Postfix instance /etc/postfix without any issues. When I change MAIL_CONFIG to point to /etc/postfix-blue in either .bash_profile or the Procmail script that handles this email, though, I get this error: sendmail: fatal: User account5(###) is not allowed to submit mail I've read the manuals on Postfix.org, searched Google, and tried the suggestions in three previous answers here on ServerFault.com: Postfix - specify interface to deliver outbound mail on Postfix user is not allowed to submit mail Postfix rejects php mails I have been careful to stop and restart Postfix after each configuration change, and tested the results. Nothing has worked. The main postfix instance happily accepts outbound email from account5. The postfix-blue instance continues to reject email from account5 with the sendmail error above. As tempting as it is to blame machine hostility, I know that I must be missing something or doing something wrong. Does anybody have any suggestions as to what it might be? Please feel free to ask for further information about my setup if you need it. =-=-=-=-=-=-=-=-=-= At the request of the responder, here are main.cf and master.cf for a) the main postfix instance ("red.example.com") and b) the FBL instance ("blue.example.us") [NOTE: All parameters not specified below were left at the default Postfix 2.6 settings] MAIN: master.cf smtp inet n - n - - smtpd main.cf myhostname = red.example.com mydomain = example.com inet_interfaces = $myhostname, localhost inet_protocols = all lmtp_host_lookup = native smtp_host_lookup = native ignore_mx_lookup_error = yes mydestination = $myhostname, localhost.$mydomain, localhost local_recipient_maps = mynetworks = 192.168.34.104/32 relay_domains = example.com, example.info, example.net, example.org, example.us relayhost = [192.168.34.102] # Separate physical server, main mailserver. relay_recipient_maps = hash:/etc/postfix/relay_recipients alias_maps = hash:/etc/aliases alias_database = hash:/etc/aliases smtpd_banner = $myhostname ESMTP $mail_name multi_instance_wrapper = ${command_directory}/postmulti -p -- multi_instance_enable = yes multi_instance_directories = /etc/postfix-green /etc/postfix-pink /etc/postfix-orange /etc/postfix-blue FBL: master.cf 184.173.119.103:25 inet n - n - - smtpd main.cf myhostname = blue.example.us mydomain = blue.example.us <= Deliberately set to subdomain only. myorigin = $mydomain inet_interfaces = $myhostname lmtp_host_lookup = native smtp_host_lookup = native ignore_mx_lookup_error = yes mydestination = $myhostname local_recipient_maps = unix:passwd.byname $alias_maps $virtual_alias_maps mynetworks = 192.168.36.51/32, 192.168.35.20/31 <= Second IP is backup MX servers relay_domains = $mydestination recipient_canonical_maps = hash:/etc/postfix-blue/canonical virtual_alias_maps = hash:/etc/postfix-fbl/virtual alias_maps = hash:/etc/aliases, hash:/etc/postfix-blue/canonical alias_maps = hash:/etc/aliases, hash:/etc/postfix-blue/canonical mailbox_command = /usr/bin/procmail -a "$EXTENSION" DEFAULT=$HOME/Mail/ MAILDIR=$HOME/Mail smtpd_banner = $myhostname ESMTP $mail_name authorized_submit_users = multi_instance_name = postfix-blue multi_instance_enable = yes

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  • Quick guide to Oracle IRM 11g: Creating your first sealed document

    - by Simon Thorpe
    Quick guide to Oracle IRM 11g indexThe previous articles in this guide have detailed how to install, configure and secure your Oracle IRM 11g service. This article walks you through the process of now creating your first context and securing a document against it. I should mention that it would be worth reviewing the following to ensure your installation is ready for that all important first document. Ensure you have correctly configured the keystore for the IRM wrapper keys. If this is not correctly configured, creating the context below will fail. Make sure the IRM server URL correctly resolves and uses the right protocol (HTTP or HTTPS) ContentsCreate the first contextInstall the Oracle IRM Desktop Seal your first document Create the first contextIn Oracle 11g there is a built in classification and rights system called the "standard rights model" which is based on 10 years of customer use cases and innovation. It is a system which enables IRM to scale massively whilst retaining the ability to balance security and usability and also separate duties by allowing contacts in the business to own classifications. The final article in this guide goes into detail on this inbuilt classification model, but for the purposes of this current article all we need to do is create at least one context to test our system out.With a new IRM server there are a set of predefined context templates and roles which again are setup in a way which reflects the most common use we've learned from our customers. We will use these out of the box configurations as they are to create the first context against which we will seal some content.First login to your Oracle IRM Management Website located at https://irm.company.com/irm_rights/. Currently the system is only configured to use the built in LDAP for users, so use the only account we have at the moment, which by default is weblogic. Once logged in switch to the Contexts tab. Click on the New Context icon () in the menu bar on the left. In the resulting dialog select the Standard context template and enter in a name for the context. Then just hit finish, the weblogic account will automatically be made the manager. You'll now see your brand new context ready for users to be assigned. Now click on the Assign Role icon () in the menu bar and in the resulting dialog search for your only user account, weblogic, and add to the list on the right. Now select a role for this user. Because we need to create a document with this user we must select contributor, as this is the only role which allows for the ability to seal. Finally hit next and then finish. We now have a context with a user that has the rights to create a document. The next step is to configure the IRM Desktop to get these rights from the server. Install the Oracle IRM Desktop Before we can seal a document we need the client software installed. Oracle IRM has a very small, lightweight client called the Oracle IRM Desktop which can be freely downloaded in 27 languages from here. Double click on the installer and click on next... Next again... And finally on install... Very easy. You may get a warning about closing Outlook, Word or another application and most of the time no reboots are required. Once it is installed you will see the IRM Desktop icon running in your tool tray, bottom right of the desktop. Seal your first document Finally the prize is within reach, creating your first sealed document. The server is running, we've got a context ready, a user assigned a role in the context but there is the simple and obvious hoop left to jump through. To seal a document we need to have the users rights cached to the local machine. For this to take place, the IRM Desktop needs to know where the Oracle IRM server is on the network so we can synchronize these rights and then be able to seal a document. The usual way for the IRM Desktop to know about the IRM server is it learns automatically when you open an existing piece of content that someone has sent you... ack. Bit of a chicken or the egg dilemma. The solution is to manually tell the IRM Desktop the location of the IRM Server and then force a synchronization of rights. Right click on the Oracle IRM Desktop icon in the system tray and select Options.... Then switch to the Servers tab in the resulting dialog. There are no servers in the list because you've never opened any content. This list is usually populated automatically but we are going to add a server manually, so click on New.... Into the dialog enter in the full URL to the IRM server. Note that this time you use the path /irm_desktop/ and not /irm_rights/. You can see an example from the image below. Click on the validate button and you'll be asked to authenticate. Enter in your weblogic username and password and also check the Remember my password check box. Click OK and the IRM Desktop will confirm a successful connection to the server. OK all the dialogs and we are ready to Synchronize this users rights to the desktop. Right click once more on the Oracle IRM Desktop icon in the system tray. Now the Synchronize menu option is available. Select this and the IRM Desktop will now talk to the IRM server, authenticate using your weblogic account and get your rights to the context we created. Because this is the first time this users has communicated with the IRM server the IRM Desktop presents a privacy policy dialog. This is a chance for the business to ask users to agree to any policy about the use of IRM before opening secured documents. In our guide we've not bothered to setup this URL so just click on the check box and hit Accept. The IRM Desktop will then talk to the server, get your rights and display a success dialog. Lets protect a documentNow we are ready to seal a piece of content. In my guide i'm going to protect a Microsoft Word document. This mean's I have to have copy of Office installed, in this guide i'm using Microsoft Office 2007. You could also seal a PDF document, you'll need to download and install Adobe Acrobat Reader. A very simple test could be to seal a GIF/JPG/PNG or piece of HTML because this is rendered using Internet Explorer. But as I say, i'm going to protect a Word document. The following example demonstrates choosing a file in Windows Explorer, there are many ways to seal a file and you can watch a few in this video.Open a copy of Windows Explorer and locate the file you wish to seal. Right click on the document and select Seal To -> Context You are now presented with the Select Context dialog. You'll now have a sealed copy of the document sat in the same location. Double click on this document and it will open, again using the credentials you've already provided. That is it, now you just need to add more users, more documents, more classifications and start exploring the different roles and experiment with different offline periods etc. You may wish to setup the server against an existing LDAP or Active Directory environment instead of using the built in WebLogic LDAP store. You can read how to use your corporate directory here. But before we finish this guide, there is one more article and arguably the most important article of all. Next I discuss the all important decision making surrounding the actually implementation of Oracle IRM inside your business. Who has rights to what? How do you map contexts to your existing business practices? It is the next article which actually ensures you deploy a successful IRM solution by looking at the business and understanding how they use your sensitive information and then configuring Oracle IRM to reflect their use.

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  • The last MVVM you'll ever need?

    - by Nuri Halperin
    As my MVC projects mature and grow, the need to have some omnipresent, ambient model properties quickly emerge. The application no longer has only one dynamic pieced of data on the page: A sidebar with a shopping cart, some news flash on the side – pretty common stuff. The rub is that a controller is invoked in context of a single intended request. The rest of the data, even though it could be just as dynamic, is expected to appear on it's own. There are many solutions to this scenario. MVVM prescribes creating elaborate objects which expose your new data as a property on some uber-object with more properties exposing the "side show" ambient data. The reason I don't love this approach is because it forces fairly acute awareness of the view, and soon enough you have many MVVM objects laying around, and views have to start doing null-checks in order to ensure you really supplied all the values before binding to them. Ick. Just as unattractive is the ViewData dictionary. It's not strongly typed, and in both this and the MVVM approach someone has to populate these properties – n'est pas? Where does that live? With MVC2, we get the formerly-futures  feature Html.RenderAction(). The feature allows you plant a line in a view, of the format: <% Html.RenderAction("SessionInterest", "Session"); %> While this syntax looks very clean, I can't help being bothered by it. MVC was touting a very strong separation of concerns, the Model taking on the role of the business logic, the controller handling route and performing minimal view-choosing operations and the views strictly focused on rendering out angled-bracket tags. The RenderAction() syntax has the view calling some controller and invoking it inline with it's runtime rendering. This – to my taste – embeds too much  knowledge of controllers into the view's code – which was allegedly forbidden.  The one way flow "Controller Receive Data –> Controller invoke Model –> Controller select view –> Controller Hand data to view" now gets a "View calls controller and gets it's own data" which is not so one-way anymore. Ick. I toyed with some other solutions a bit, including some base controllers, special view classes etc. My current favorite though is making use of the ExpandoObject and dynamic features with C# 4.0. If you follow Phil Haack or read a bit from David Heyden you can see the general picture emerging. The game changer is that using the new dynamic syntax, one can sprout properties on an object and make use of them in the view. Well that beats having a bunch of uni-purpose MVVM's any day! Rather than statically exposed properties, we'll just use the capability of adding members at runtime. Armed with new ideas and syntax, I went to work: First, I created a factory method to enrich the focuse object: public static class ModelExtension { public static dynamic Decorate(this Controller controller, object mainValue) { dynamic result = new ExpandoObject(); result.Value = mainValue; result.SessionInterest = CodeCampBL.SessoinInterest(); result.TagUsage = CodeCampBL.TagUsage(); return result; } } This gives me a nice fluent way to have the controller add the rest of the ambient "side show" items (SessionInterest, TagUsage in this demo) and expose them all as the Model: public ActionResult Index() { var data = SyndicationBL.Refresh(TWEET_SOURCE_URL); dynamic result = this.Decorate(data); return View(result); } So now what remains is that my view knows to expect a dynamic object (rather than statically typed) so that the ASP.NET page compiler won't barf: <%@ Page Language="C#" Title="Ambient Demo" MasterPageFile="~/Views/Shared/Ambient.Master" Inherits="System.Web.Mvc.ViewPage<dynamic>" %> Notice the generic ViewPage<dynamic>. It doesn't work otherwise. In the page itself, Model.Value property contains the main data returned from the controller. The nice thing about this, is that the master page (Ambient.Master) also inherits from the generic ViewMasterPage<dynamic>. So rather than the page worrying about all this ambient stuff, the side bars and panels for ambient data all reside in a master page, and can be rendered using the RenderPartial() syntax: <% Html.RenderPartial("TagCloud", Model.SessionInterest as Dictionary<string, int>); %> Note here that a cast is necessary. This is because although dynamic is magic, it can't figure out what type this property is, and wants you to give it a type so its binder can figure out the right property to bind to at runtime. I use as, you can cast if you like. So there we go – no violation of MVC, no explosion of MVVM models and voila – right? Well, I could not let this go without a tweak or two more. The first thing to improve, is that some views may not need all the properties. In that case, it would be a waste of resources to populate every property. The solution to this is simple: rather than exposing properties, I change d the factory method to expose lambdas - Func<T> really. So only if and when a view accesses a member of the dynamic object does it load the data. public static class ModelExtension { // take two.. lazy loading! public static dynamic LazyDecorate(this Controller c, object mainValue) { dynamic result = new ExpandoObject(); result.Value = mainValue; result.SessionInterest = new Func<Dictionary<string, int>>(() => CodeCampBL.SessoinInterest()); result.TagUsage = new Func<Dictionary<string, int>>(() => CodeCampBL.TagUsage()); return result; } } Now that lazy loading is in place, there's really no reason not to hook up all and any possible ambient property. Go nuts! Add them all in – they won't get invoked unless used. This now requires changing the signature of usage on the ambient properties methods –adding some parenthesis to the master view: <% Html.RenderPartial("TagCloud", Model.SessionInterest() as Dictionary<string, int>); %> And, of course, the controller needs to call LazyDecorate() rather than the old Decorate(). The final touch is to introduce a convenience method to the my Controller class , so that the tedium of calling Decorate() everywhere goes away. This is done quite simply by adding a bunch of methods, matching View(object), View(string,object) signatures of the Controller class: public ActionResult Index() { var data = SyndicationBL.Refresh(TWEET_SOURCE_URL); return AmbientView(data); } //these methods can reside in a base controller for the solution: public ViewResult AmbientView(dynamic data) { dynamic result = ModelExtension.LazyDecorate(this, data); return View(result); } public ViewResult AmbientView(string viewName, dynamic data) { dynamic result = ModelExtension.LazyDecorate(this, data); return View(viewName, result); } The call to AmbientView now replaces any call the View() that requires the ambient data. DRY sattisfied, lazy loading and no need to replace core pieces of the MVC pipeline. I call this a good MVC day. Enjoy!

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  • DBCC CHECKDB on VVLDB and latches (Or: My Pain is Your Gain)

    - by Argenis
      Does your CHECKDB hurt, Argenis? There is a classic blog series by Paul Randal [blog|twitter] called “CHECKDB From Every Angle” which is pretty much mandatory reading for anybody who’s even remotely considering going for the MCM certification, or its replacement (the Microsoft Certified Solutions Master: Data Platform – makes my fingers hurt just from typing it). Of particular interest is the post “Consistency Options for a VLDB” – on it, Paul provides solid, timeless advice (I use the word “timeless” because it was written in 2007, and it all applies today!) on how to perform checks on very large databases. Well, here I was trying to figure out how to make CHECKDB run faster on a restored copy of one of our databases, which happens to exceed 7TB in size. The whole thing was taking several days on multiple systems, regardless of the storage used – SAS, SATA or even SSD…and I actually didn’t pay much attention to how long it was taking, or even bothered to look at the reasons why - as long as it was finishing okay and found no consistency errors. Yes – I know. That was a huge mistake, as corruption found in a database several days after taking place could only allow for further spread of the corruption – and potentially large data loss. In the last two weeks I increased my attention towards this problem, as we noticed that CHECKDB was taking EVEN LONGER on brand new all-flash storage in the SAN! I couldn’t really explain it, and were almost ready to blame the storage vendor. The vendor told us that they could initially see the server driving decent I/O – around 450Mb/sec, and then it would settle at a very slow rate of 10Mb/sec or so. “Hum”, I thought – “CHECKDB is just not pushing the I/O subsystem hard enough”. Perfmon confirmed the vendor’s observations. Dreaded @BlobEater What was CHECKDB doing all the time while doing so little I/O? Eating Blobs. It turns out that CHECKDB was taking an extremely long time on one of our frankentables, which happens to be have 35 billion rows (yup, with a b) and sucks up several terabytes of space in the database. We do have a project ongoing to purge/split/partition this table, so it’s just a matter of time before we deal with it. But the reality today is that CHECKDB is coming to a screeching halt in performance when dealing with this particular table. Checking sys.dm_os_waiting_tasks and sys.dm_os_latch_stats showed that LATCH_EX (DBCC_OBJECT_METADATA) was by far the top wait type. I remembered hearing recently about that wait from another post that Paul Randal made, but that was related to computed-column indexes, and in fact, Paul himself reminded me of his article via twitter. But alas, our pathologic table had no non-clustered indexes on computed columns. I knew that latches are used by the database engine to do internal synchronization – but how could I help speed this up? After all, this is stuff that doesn’t have a lot of knobs to tweak. (There’s a fantastic level 500 talk by Bob Ward from Microsoft CSS [blog|twitter] called “Inside SQL Server Latches” given at PASS 2010 – and you can check it out here. DISCLAIMER: I assume no responsibility for any brain melting that might ensue from watching Bob’s talk!) Failed Hypotheses Earlier on this week I flew down to Palo Alto, CA, to visit our Headquarters – and after having a great time with my Monkey peers, I was relaxing on the plane back to Seattle watching a great talk by SQL Server MVP and fellow MCM Maciej Pilecki [twitter] called “Masterclass: A Day in the Life of a Database Transaction” where he discusses many different topics related to transaction management inside SQL Server. Very good stuff, and when I got home it was a little late – that slow DBCC CHECKDB that I had been dealing with was way in the back of my head. As I was looking at the problem at hand earlier on this week, I thought “How about I set the database to read-only?” I remembered one of the things Maciej had (jokingly) said in his talk: “if you don’t want locking and blocking, set the database to read-only” (or something to that effect, pardon my loose memory). I immediately killed the CHECKDB which had been running painfully for days, and set the database to read-only mode. Then I ran DBCC CHECKDB against it. It started going really fast (even a bit faster than before), and then throttled down again to around 10Mb/sec. All sorts of expletives went through my head at the time. Sure enough, the same latching scenario was present. Oh well. I even spent some time trying to figure out if NUMA was hurting performance. Folks on Twitter made suggestions in this regard (thanks, Lonny! [twitter]) …Eureka? This past Friday I was still scratching my head about the whole thing; I was ready to start profiling with XPERF to see if I could figure out which part of the engine was to blame and then get Microsoft to look at the evidence. After getting a bunch of good news I’ll blog about separately, I sat down for a figurative smack down with CHECKDB before the weekend. And then the light bulb went on. A sparse column. I thought that I couldn’t possibly be experiencing the same scenario that Paul blogged about back in March showing extreme latching with non-clustered indexes on computed columns. Did I even have a non-clustered index on my sparse column? As it turns out, I did. I had one filtered non-clustered index – with the sparse column as the index key (and only column). To prove that this was the problem, I went and setup a test. Yup, that'll do it The repro is very simple for this issue: I tested it on the latest public builds of SQL Server 2008 R2 SP2 (CU6) and SQL Server 2012 SP1 (CU4). First, create a test database and a test table, which only needs to contain a sparse column: CREATE DATABASE SparseColTest; GO USE SparseColTest; GO CREATE TABLE testTable (testCol smalldatetime SPARSE NULL); GO INSERT INTO testTable (testCol) VALUES (NULL); GO 1000000 That’s 1 million rows, and even though you’re inserting NULLs, that’s going to take a while. In my laptop, it took 3 minutes and 31 seconds. Next, we run DBCC CHECKDB against the database: DBCC CHECKDB('SparseColTest') WITH NO_INFOMSGS, ALL_ERRORMSGS; This runs extremely fast, as least on my test rig – 198 milliseconds. Now let’s create a filtered non-clustered index on the sparse column: CREATE NONCLUSTERED INDEX [badBadIndex] ON testTable (testCol) WHERE testCol IS NOT NULL; With the index in place now, let’s run DBCC CHECKDB one more time: DBCC CHECKDB('SparseColTest') WITH NO_INFOMSGS, ALL_ERRORMSGS; In my test system this statement completed in 11433 milliseconds. 11.43 full seconds. Quite the jump from 198 milliseconds. I went ahead and dropped the filtered non-clustered indexes on the restored copy of our production database, and ran CHECKDB against that. We went down from 7+ days to 19 hours and 20 minutes. Cue the “Argenis is not impressed” meme, please, Mr. LaRock. My pain is your gain, folks. Go check to see if you have any of such indexes – they’re likely causing your consistency checks to run very, very slow. Happy CHECKDBing, -Argenis ps: I plan to file a Connect item for this issue – I consider it a pretty serious bug in the engine. After all, filtered indexes were invented BECAUSE of the sparse column feature – and it makes a lot of sense to use them together. Watch this space and my twitter timeline for a link.

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