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  • Cannot get to configure Kerberos for Reporting Services

    - by Ucodia
    Context I am trying to configure Kerberos in the domain for double-hop authentication. So here are the machines and their respective roles: client01: Windows 7 as client dc01: Windows Server 2008 R2 as domain controller and dns server01: Windows Server 2008 R2 as reporting server (native mode) server02: Windows Server 2008 R2 as SQL Server database engine I want my client01 to connect to server01 and configure a data source that is located on server02 using Intergrated Security. So as NTLM cannot push credentials that far, I need to setup Kerberos to enable double-hop authentication. The reporting service is runned by the Network Service service account and is configured only with the RSWindowsNegotiate options for authentication. Issue I cannot get to pass my client01 credential to server02 when configuring the data source on server01. Therefore I get the error: Login failed for user 'NT AUTHORITY\ANONYMOUS LOGON'. So I went on dc01 and delegated full trust for any service to server01 but it not fixed the problem. I want to notice that I did not configured any SPNs for server01 because Reporting Service is runned by Network Service and from what I read on the Internet, when Reporting Services is going up with Network Service, SPNs are automatically registered. My problem is that even if that I want to configure SPNs manually, I do not know where I have to set them up. On dc01 or on server01? So I went a bit further on the issue and tried to trace this problem. From my understanding of Kerberos, this is what should happen on the network when I try to connect the data source: client01 ---- AS_REQ ---> dc01 <--- AS_REP ---- client01 ---- TGS_REQ ---> dc01 <--- TGS_REP ---- client01 ---- AP_REQ ---> server01 <--- AP_REP ---- server01 ---- TGS_REQ ---> dc01 <--- TGS_REP ---- server01 ---- AP_REQ ---> server02 <--- AP_REP ---- So captured my local network with Wireshark, but whenever I try to configure my data source from client01 on server01 to pass my credentials to server02, my client never sends a AS_REQ or TGS_REQ to the KDC on dc01. Questions So does anyone can tell me if I should configure the SPNs and on which machine does it have to be configured? Also why client01 never request for a TGT or a TGS to my KDC. Do you think there is something going wrong with the DC role of dc01?

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  • unable to connect site to different port

    - by JohnMerlino
    I have a domain was registered at godaddy named http://mysite.com/. I logged into godaddy and I went to All Products Domains Domain Management. I clicked on the appropriate domain and it took me to the Domain Details page. I clicked Launch under DNS Manager and it took me to the Zone File Editor. I noticed that notify.mysite.com was pointing to an IP address pointing to a dead server, so I switched it to an operating server. Then I pinged the domain to see where it was pointing to and it was correctly pointing to the working server. So I copied the default configuration under sites-available: sudo cp default notify.mysite.com. And then I made some edits to it to have it point to a different document root to serve files at a different port: Listen 1740 Listen 64.135.xx.xxx:1740 #I also tried this as well: NameVirtualHost 64.135.xx.xxx:1740 <VirtualHost 64.135.xx.xxx:1740> ServerAdmin [email protected] ServerName notify.mysite.com DocumentRoot /var/www/test/public <Directory /var/www/test/public> Order allow,deny allow from all </Directory> ErrorLog ${APACHE_LOG_DIR}/error.log LogLevel warn CustomLog ${APACHE_LOG_DIR}/access.log combined </VirtualHost> Then I enabled the virtual host. Then I went to the document root and added an index.html file with some text in it. Then I restarted apache. The restart gave no errors. Then I type the correct domain in URL: http://notify.mysite.com:1740/ and I get: Oops! Google Chrome could not connect to notify.mysite.com:1740 Somehow it took out all my other sites. Now even the ones that were responding on port 80 are no longe responding, even though I did not touch the virtual hosts for them. I get this message now: Oops! Google Chrome could not connect to mysite.com However, ping responds: ping mysite.com PING mysite.com (64.135.12.134): 56 data bytes 64 bytes from 64.135.12.134: icmp_seq=0 ttl=49 time=20.839 ms 64 bytes from 64.135.12.134: icmp_seq=1 ttl=49 time=20.489 ms The result of telnet: $ telnet guarddoggps.com 80 Trying 64.135.12.134... telnet: connect to address 64.135.12.134: Connection refused telnet: Unable to connect to remote host

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  • 2 servers, high availability and faster response

    - by user17886
    I recently bought a second webserver because I worry about hardware failure of my old server. Now that I have that second server I wish to do a little more then just have one server standby and replicate all day. As long as it's there I might as well get some advantage our of it ! I have a website powered by ubuntu 12.04, nginx, php-fpm, apc, mysql (5.5) and couchdb. Im currently testing configurations where i can achieve failover AND make good use of the extra harware for faster responses / distributed load. The setup I am testing nowinvolves heartbeat for ip failover and two identical servers. Of the two servers only one has a public ip adress. If one server crashes the other server takes over the public ip adress. On an incoming request nginx forwards the request tot php-fpm to either server a of server b (50/50 if both servers are alive). Once the request has been send to php-fpm both servers look at localhost for the mysql server. I use master-master mysql replication for this. The file system is synced with lsyncd. This works pretty well but Im reading it's discouraged by the (mysql) community. Another option I could think of is to use one server as a mysql master and one server as a web/php server. The servers would still sync their filesystem, would still run the same duplicate software (nginx,mysql) but master slave mysql replication could be used. As long as bother servers are alive I could just prefer nginx to listen to ip a and mysql to ip b. If one server is down, the other server could take over the task of the other server, simply by ip switching. But im completely new at this so I would greatly value your expert advice. Is either of the two setups any good ? If you have any thoughts on this please let me know ! PS, virtualisation, hosting on different locations or active/passive setups are not solutions im looking for. I find virtual server either too slow or too expensive. I already have a passive failover on another location. But in case of a crash I found the site was still unreachable for too long due to dns caching.

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  • Internet Pings but Does Not Load

    - by t3techcom18
    From what I've been seeing and been doing my research for the past two days, many people have been having the same issues throughout the years, however, this is the first time I've encountered this issue and many of the specific workarounds or fixes have not worked for me. I've been trying to work through this for 24 hours straight now, but to no avail so many thanks to those that can help. On Monday night, got home from work; surfing the internet for half an hour, everything was fine as always. Just after half an hour, my Internet got very sluggish and then it died completely. I thought it might have been the an update I just put through in terms of Windows Update that said was a critical update for MSE, as the same thing happened a few years ago. I did a System Restore to two different dates that were in the past two weeks, nothing. Uninstalled MSE and disabled Windows Defender and the Windows Firewall: Nothing. Reset IE Options, Reset Winsock, Dumping DNS, many of the other command prompt screens to reset items: Nothing. Reset the modem: Nothing. What DID work, however, was a ping test to Yahoo. The ping test worked, saying all four packets was recieved, yet nothing else popped up. LAN and CenturyLink said everything worked on their end and that everything was connected properly, as well as the speeds working fine. CenturyLink said in their notes that they thought Port 80 was blocked. I went and put in the Firewall to allow Port 80 but it didn't make any difference whatsoever. I remembered I had a spare modem laying around and I switched them up, both modem and the cords - nothing. I then hooked it up to my netbook to see if that would work, as it usually does - connection didn't work there either. Like I said, it's been about 24 hours now and this is increasingly frustrating, as I've tried all solutions (While browsing through 10 search results pages on my phone) suggested and still nothing. Any suggestions and tricks would be greatly appreciated! Here's my specs: Windows 7 32-bit Home Premium Intel Core 2 Duo 3.14 Ghz 4 GB Kingston DDR2 RAM eVGA nForce 750i SLI eVGA GeForce GTX 560 Ti FPB ISP: CenturyLink No router Modem: CenturyLink 660 Series Hardwired connection PLEASE NOTE: This is the only computer I have (Like I said, the netbook solution didn't work), so downloading programs and such is not an option til I get to other computers somewhere else, like right now. Unless someone knows of a way of copying/pasting a file in Windows and then transferring said info to an Android smartphone, this is gunna take a while haha. Patience is requested.

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  • WPF Blurry Images - Bitmap Class

    - by Luke
    I am using the following sample at http://blogs.msdn.com/dwayneneed/archive/2007/10/05/blurry-bitmaps.aspx within VB.NET. The code is shown below. I am having a problem when my application loads the CPU is pegging 50-70%. I have determined that the problem is with the Bitmap class. The OnLayoutUpdated() method is calling the InvalidateVisual() continously. This is because some points are not returning as equal but rather, Point(0.0,-0.5) Can anyone see any bugs within this code or know a better implmentation for pixel snapping a Bitmap image so it is not blurry? p.s. The sample code was in C#, however I believe that it was converted correctly. Imports System Imports System.Collections.Generic Imports System.Windows Imports System.Windows.Media Imports System.Windows.Media.Imaging Class Bitmap Inherits FrameworkElement ' Use FrameworkElement instead of UIElement so Data Binding works as expected Private _sourceDownloaded As EventHandler Private _sourceFailed As EventHandler(Of ExceptionEventArgs) Private _pixelOffset As Windows.Point Public Sub New() _sourceDownloaded = New EventHandler(AddressOf OnSourceDownloaded) _sourceFailed = New EventHandler(Of ExceptionEventArgs)(AddressOf OnSourceFailed) AddHandler LayoutUpdated, AddressOf OnLayoutUpdated End Sub Public Shared ReadOnly SourceProperty As DependencyProperty = DependencyProperty.Register("Source", GetType(BitmapSource), GetType(Bitmap), New FrameworkPropertyMetadata(Nothing, FrameworkPropertyMetadataOptions.AffectsRender Or FrameworkPropertyMetadataOptions.AffectsMeasure, New PropertyChangedCallback(AddressOf Bitmap.OnSourceChanged))) Public Property Source() As BitmapSource Get Return DirectCast(GetValue(SourceProperty), BitmapSource) End Get Set(ByVal value As BitmapSource) SetValue(SourceProperty, value) End Set End Property Public Shared Function FindParentWindow(ByVal child As DependencyObject) As Window Dim parent As DependencyObject = VisualTreeHelper.GetParent(child) 'Check if this is the end of the tree If parent Is Nothing Then Return Nothing End If Dim parentWindow As Window = TryCast(parent, Window) If parentWindow IsNot Nothing Then Return parentWindow Else ' Use recursion until it reaches a Window Return FindParentWindow(parent) End If End Function Public Event BitmapFailed As EventHandler(Of ExceptionEventArgs) ' Return our measure size to be the size needed to display the bitmap pixels. ' ' Use MeasureOverride instead of MeasureCore so Data Binding works as expected. ' Protected Overloads Overrides Function MeasureCore(ByVal availableSize As Size) As Size Protected Overloads Overrides Function MeasureOverride(ByVal availableSize As Size) As Size Dim measureSize As New Size() Dim bitmapSource As BitmapSource = Source If bitmapSource IsNot Nothing Then Dim ps As PresentationSource = PresentationSource.FromVisual(Me) If Me.VisualParent IsNot Nothing Then Dim window As Window = window.GetWindow(Me.VisualParent) If window IsNot Nothing Then ps = PresentationSource.FromVisual(window.GetWindow(Me.VisualParent)) ElseIf FindParentWindow(Me) IsNot Nothing Then ps = PresentationSource.FromVisual(FindParentWindow(Me)) End If End If ' If ps IsNot Nothing Then Dim fromDevice As Matrix = ps.CompositionTarget.TransformFromDevice Dim pixelSize As New Vector(bitmapSource.PixelWidth, bitmapSource.PixelHeight) Dim measureSizeV As Vector = fromDevice.Transform(pixelSize) measureSize = New Size(measureSizeV.X, measureSizeV.Y) Else measureSize = New Size(bitmapSource.PixelWidth, bitmapSource.PixelHeight) End If End If Return measureSize End Function Protected Overloads Overrides Sub OnRender(ByVal dc As DrawingContext) Dim bitmapSource As BitmapSource = Me.Source If bitmapSource IsNot Nothing Then _pixelOffset = GetPixelOffset() ' Render the bitmap offset by the needed amount to align to pixels. dc.DrawImage(bitmapSource, New Rect(_pixelOffset, DesiredSize)) End If End Sub Private Shared Sub OnSourceChanged(ByVal d As DependencyObject, ByVal e As DependencyPropertyChangedEventArgs) Dim bitmap As Bitmap = DirectCast(d, Bitmap) Dim oldValue As BitmapSource = DirectCast(e.OldValue, BitmapSource) Dim newValue As BitmapSource = DirectCast(e.NewValue, BitmapSource) If ((oldValue IsNot Nothing) AndAlso (bitmap._sourceDownloaded IsNot Nothing)) AndAlso (Not oldValue.IsFrozen AndAlso (TypeOf oldValue Is BitmapSource)) Then RemoveHandler DirectCast(oldValue, BitmapSource).DownloadCompleted, bitmap._sourceDownloaded RemoveHandler DirectCast(oldValue, BitmapSource).DownloadFailed, bitmap._sourceFailed ' ((BitmapSource)newValue).DecodeFailed -= bitmap._sourceFailed; // 3.5 End If If ((newValue IsNot Nothing) AndAlso (TypeOf newValue Is BitmapSource)) AndAlso Not newValue.IsFrozen Then AddHandler DirectCast(newValue, BitmapSource).DownloadCompleted, bitmap._sourceDownloaded AddHandler DirectCast(newValue, BitmapSource).DownloadFailed, bitmap._sourceFailed ' ((BitmapSource)newValue).DecodeFailed += bitmap._sourceFailed; // 3.5 End If End Sub Private Sub OnSourceDownloaded(ByVal sender As Object, ByVal e As EventArgs) InvalidateMeasure() InvalidateVisual() End Sub Private Sub OnSourceFailed(ByVal sender As Object, ByVal e As ExceptionEventArgs) Source = Nothing ' setting a local value seems scetchy... RaiseEvent BitmapFailed(Me, e) End Sub Private Sub OnLayoutUpdated(ByVal sender As Object, ByVal e As EventArgs) ' This event just means that layout happened somewhere. However, this is ' what we need since layout anywhere could affect our pixel positioning. Dim pixelOffset As Windows.Point = GetPixelOffset() If Not AreClose(pixelOffset, _pixelOffset) Then InvalidateVisual() End If End Sub ' Gets the matrix that will convert a Windows.Point from "above" the ' coordinate space of a visual into the the coordinate space ' "below" the visual. Private Function GetVisualTransform(ByVal v As Visual) As Matrix If v IsNot Nothing Then Dim m As Matrix = Matrix.Identity Dim transform As Transform = VisualTreeHelper.GetTransform(v) If transform IsNot Nothing Then Dim cm As Matrix = transform.Value m = Matrix.Multiply(m, cm) End If Dim offset As Vector = VisualTreeHelper.GetOffset(v) m.Translate(offset.X, offset.Y) Return m End If Return Matrix.Identity End Function Private Function TryApplyVisualTransform(ByVal Point As Windows.Point, ByVal v As Visual, ByVal inverse As Boolean, ByVal throwOnError As Boolean, ByRef success As Boolean) As Windows.Point success = True If v IsNot Nothing Then Dim visualTransform As Matrix = GetVisualTransform(v) If inverse Then If Not throwOnError AndAlso Not visualTransform.HasInverse Then success = False Return New Windows.Point(0, 0) End If visualTransform.Invert() End If Point = visualTransform.Transform(Point) End If Return Point End Function Private Function ApplyVisualTransform(ByVal Point As Windows.Point, ByVal v As Visual, ByVal inverse As Boolean) As Windows.Point Dim success As Boolean = True Return TryApplyVisualTransform(Point, v, inverse, True, success) End Function Private Function GetPixelOffset() As Windows.Point Dim pixelOffset As New Windows.Point() Dim ps As PresentationSource = PresentationSource.FromVisual(Me) If ps IsNot Nothing Then Dim rootVisual As Visual = ps.RootVisual ' Transform (0,0) from this element up to pixels. pixelOffset = Me.TransformToAncestor(rootVisual).Transform(pixelOffset) pixelOffset = ApplyVisualTransform(pixelOffset, rootVisual, False) pixelOffset = ps.CompositionTarget.TransformToDevice.Transform(pixelOffset) ' Round the origin to the nearest whole pixel. pixelOffset.X = Math.Round(pixelOffset.X) pixelOffset.Y = Math.Round(pixelOffset.Y) ' Transform the whole-pixel back to this element. pixelOffset = ps.CompositionTarget.TransformFromDevice.Transform(pixelOffset) pixelOffset = ApplyVisualTransform(pixelOffset, rootVisual, True) pixelOffset = rootVisual.TransformToDescendant(Me).Transform(pixelOffset) End If Return pixelOffset End Function Private Function AreClose(ByVal Point1 As Windows.Point, ByVal Point2 As Windows.Point) As Boolean Return AreClose(Point1.X, Point2.X) AndAlso AreClose(Point1.Y, Point2.Y) End Function Private Function AreClose(ByVal value1 As Double, ByVal value2 As Double) As Boolean If value1 = value2 Then Return True End If Dim delta As Double = value1 - value2 Return ((delta < 0.00000153) AndAlso (delta > -0.00000153)) End Function End Class

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  • Cant correctly install Lazarus

    - by user206316
    I have a little problem with installing and running Lazarus. I just upgrade ubuntu from 13.04 to 13.10. When i had 13.04, i could install lazarus without any problems, but in 13.10 lazarus magicaly dissapeared, and when i tried install it from ubuntu software center, it said something like in my software resources lazarus-ide-0.9.30.4 doesnt exist. After some research on net i tried delete all files from earlier installations, download deb packages from sourceforge and install them, but when i want to instal fpc-src, error shows up with output: (Reading database ... 100% (Reading database ... 239063 files and directories currently installed.) Unpacking fpc-src (from .../Stiahnut/Lazarus/fpc-src.deb) ... dpkg: error processing /home/richi/Stiahnut/Lazarus/fpc-src.deb (--install): trying to overwrite '/usr/share/fpcsrc/2.6.2/rtl/nativent/tthread.inc', which is also in package fpc-source-2.6.2 2.6.2-5 dpkg-deb (subprocess): decompressing archive member: internal gzip write error: Broken pipe dpkg-deb: error: subprocess <decompress> returned error exit status 2 dpkg-deb (subprocess): cannot copy archive member from '/home/richi/Stiahnut/Lazarus/fpc-src.deb' to decompressor pipe: failed to write (Broken pipe) when i started lazarus, it of course tell me that it cant find fpc compier and fpc sources. So, please, i really need program for school and i dont wanna reinstall os anymore or something like that :( (Ubuntu 13.10 64bit) P.S: im not skilled in linux so if u know some commands to fix it just write them for copy and paste :) P.P.S:Sorry for bad English, im Slovak xD P.P.P.S: Thank so much for any answers update: output from sudo dpkg -l | grep "^rc" richi@Richi-Ubuntu:~/lazarus1.0.12$ sudo dpkg -l | grep "^rc" rc account-plugin-generic-oauth 0.10bzr13.03.26-0ubuntu1.1 amd64 GNOME Control Center account plugin for single signon - generic OAuth rc appmenu-gtk:amd64 12.10.3daily13.04.03-0ubuntu1 amd64 Export GTK menus over DBus rc appmenu-gtk3:amd64 12.10.3daily13.04.03-0ubuntu1 amd64 Export GTK menus over DBus rc fp-compiler-2.6.0 2.6.0-9 amd64 Free Pascal - compiler rc fp-utils-2.6.0 2.6.0-9 amd64 Free Pascal - utilities rc lazarus-ide-0.9.30.4 0.9.30.4-4 amd64 IDE for Free Pascal - common IDE files rc lazarus-ide-1.0.10 1.0.10+dfsg-1 amd64 IDE for Free Pascal - common IDE files rc lcl-utils-0.9.30.4 0.9.30.4-4 amd64 Lazarus Components Library - command line build tools rc lcl-utils-1.0.10 1.0.10+dfsg-1 amd64 Lazarus Components Library - command line build tools rc libbamf3-1:amd64 0.4.0daily13.06.19~13.04-0ubuntu1 amd64 Window matching library - shared library rc libboost-filesystem1.49.0 1.49.0-4 amd64 filesystem operations (portable paths, iteration over directories, etc) in C++ rc libboost-signals1.49.0 1.49.0-4 amd64 managed signals and slots library for C++ rc libboost-system1.49.0 1.49.0-4 amd64 Operating system (e.g. diagnostics support) library rc libboost-thread1.49.0 1.49.0-4 amd64 portable C++ multi-threading rc libbrlapi0.5:amd64 4.4-8ubuntu4 amd64 braille display access via BRLTTY - shared library rc libcamel-1.2-40 3.6.4-0ubuntu1.1 amd64 Evolution MIME message handling library rc libcolumbus0-0 0.4.0daily13.04.16~13.04-0ubuntu1 amd64 error tolerant matching engine - shared library rc libdns95 1:9.9.2.dfsg.P1-2ubuntu2.1 amd64 DNS Shared Library used by BIND rc libdvbpsi7 0.2.2-1 amd64 library for MPEG TS and DVB PSI tables decoding and generating rc libebackend-1.2-5 3.6.4-0ubuntu1.1 amd64 Utility library for evolution data servers rc libedata-book-1.2-15 3.6.4-0ubuntu1.1 amd64 Backend library for evolution address books rc libedata-cal-1.2-18 3.6.4-0ubuntu1.1 amd64 Backend library for evolution calendars rc libgc1c3:amd64 1:7.2d-0ubuntu5 amd64 conservative garbage collector for C and C++ rc libgd2-xpm:amd64 2.0.36~rc1~dfsg-6.1ubuntu1 amd64 GD Graphics Library version 2 rc libgd2-xpm:i386 2.0.36~rc1~dfsg-6.1ubuntu1 i386 GD Graphics Library version 2 rc libgnome-desktop-3-4 3.6.3-0ubuntu1 amd64 Utility library for loading .desktop files - runtime files rc libgphoto2-2:amd64 2.4.14-2 amd64 gphoto2 digital camera library rc libgphoto2-2:i386 2.4.14-2 i386 gphoto2 digital camera library rc libgphoto2-port0:amd64 2.4.14-2 amd64 gphoto2 digital camera port library rc libgphoto2-port0:i386 2.4.14-2 i386 gphoto2 digital camera port library rc libgtksourceview-3.0-0:amd64 3.6.3-0ubuntu1 amd64 shared libraries for the GTK+ syntax highlighting widget rc libgweather-3-1 3.6.2-0ubuntu1 amd64 GWeather shared library rc libharfbuzz0:amd64 0.9.13-1 amd64 OpenType text shaping engine rc libibus-1.0-0:amd64 1.4.2-0ubuntu2 amd64 Intelligent Input Bus - shared library rc libical0 0.48-2 amd64 iCalendar library implementation in C (runtime) rc libimobiledevice3 1.1.4-1ubuntu6.2 amd64 Library for communicating with the iPhone and iPod Touch rc libisc92 1:9.9.2.dfsg.P1-2ubuntu2.1 amd64 ISC Shared Library used by BIND rc libkms1:amd64 2.4.46-1 amd64 Userspace interface to kernel DRM buffer management rc libllvm3.2:i386 1:3.2repack-7ubuntu1 i386 Low-Level Virtual Machine (LLVM), runtime library rc libmikmod2:amd64 3.1.12-5 amd64 Portable sound library rc libpackagekit-glib2-14:amd64 0.7.6-3ubuntu1 amd64 Library for accessing PackageKit using GLib rc libpoppler28:amd64 0.20.5-1ubuntu3 amd64 PDF rendering library rc libraw5:amd64 0.14.7-0ubuntu1.13.04.2 amd64 raw image decoder library rc librhythmbox-core6 2.98-0ubuntu5 amd64 support library for the rhythmbox music player rc libsdl-mixer1.2:amd64 1.2.12-7ubuntu1 amd64 Mixer library for Simple DirectMedia Layer 1.2, libraries rc libsnmp15 5.4.3~dfsg-2.7ubuntu1 amd64 SNMP (Simple Network Management Protocol) library rc libsyncdaemon-1.0-1 4.2.0-0ubuntu1 amd64 Ubuntu One synchronization daemon library rc libunity-core-6.0-5 7.0.0daily13.06.19~13.04-0ubuntu1 amd64 Core library for the Unity interface. rc libusb-0.1-4:i386 2:0.1.12-23.2ubuntu1 i386 userspace USB programming library rc libwayland0:amd64 1.0.5-0ubuntu1 amd64 wayland compositor infrastructure - shared libraries rc linux-image-3.8.0-19-generic 3.8.0-19.30 amd64 Linux kernel image for version 3.8.0 on 64 bit x86 SMP rc linux-image-3.8.0-31-generic 3.8.0-31.46 amd64 Linux kernel image for version 3.8.0 on 64 bit x86 SMP rc linux-image-extra-3.8.0-19-generic 3.8.0-19.30 amd64 Linux kernel image for version 3.8.0 on 64 bit x86 SMP rc linux-image-extra-3.8.0-31-generic 3.8.0-31.46 amd64 Linux kernel image for version 3.8.0 on 64 bit x86 SMP rc screen-resolution-extra 0.15ubuntu1 all Extension for the GNOME screen resolution applet rc unity-common 7.0.0daily13.06.19~13.04-0ubuntu1 all Common files for the Unity interface.

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  • Lessons from a SAN Failure

    - by Bill Graziano
    At 1:10AM Sunday morning the main SAN at one of my clients suffered a “partial” failure.  Partial means that the SAN was still online and functioning but the LUNs attached to our two main SQL Servers “failed”.  Failed means that SQL Server wouldn’t start and the MDF and LDF files mostly showed a zero file size.  But they were online and responding and most other LUNs were available.  I’m not sure how SANs know to fail at 1AM on a Saturday night but they seem to.  From a personal standpoint this worked out poorly: I was out with friends and after more than a few drinks.  From a work standpoint this was about the best time to fail you could imagine.  Everything was running well before Monday morning.  But it was a long, long Sunday.  I started tipsy, got tired and ended up hung over later in the day. Note to self: Try not to go out drinking right before the SAN fails. This caught us at an interesting time.  We’re in the process of migrating to an entirely new set of servers so some things were partially moved.  This made it difficult to follow our procedures as cleanly as we’d like.  The benefit was that we had much better documentation of everything on the server.  I would encourage everyone to really think through the process of implementing your DR plan and document as much as possible.  Following a checklist is much easier than trying to remember at night under pressure in a hurry after a few drinks. I had a series of estimates on how long things would take.  They were accurate for any single server failure.  They weren’t accurate for a SAN failure that took two servers down.  This wasn’t bad but we should have communicated better. Don’t forget how many things are outside the database.  Logins, linked servers, DTS packages (yikes!), jobs, service broker, DTC (especially DTC), database triggers and any objects in the master database are all things you need backed up.  We’d done a decent job on this and didn’t find significant problems here.  That said this still took a lot of time.  There were many annoyances as a result of this.  Small settings like a login’s default database had a big impact on whether an application could run.  This is probably the single biggest area of concern when looking to recreate a server.  I’d encourage everyone to go through every single node of SSMS and look for user created objects or settings outside the database. Script out your logins with the proper SID and already encrypted passwords and keep it updated.  This makes life so much easier.  I used an approach based on KB246133 that worked well.  I’ll get my scripts posted over the next few days. The disaster can cause your DR process to fail in unexpected ways.  We have a job that scripts out all logins and role memberships and writes it to a file.  This runs on the DR server and pulls from the production server.  Upon opening the file I found that the contents were a “server not found” error.  Fortunately we had other copies and didn’t need to try and restore the master database.  This now runs on the production server and pushes the script to the DR site.  Soon we’ll get it pushed to our version control software. One of the biggest challenges is keeping your DR resources up to date.  Any server change (new linked server, new SQL Server Agent job, etc.) means that your DR plan (and scripts) is out of date.  It helps to automate the generation of these resources if possible. Take time now to test your database restore process.  We test ours quarterly.  If you have a large database I’d also encourage you to invest in a compressed backup solution.  Restoring backups was the single larger consumer of time during our recovery. And yes, there’s a database mirroring solution planned in our new architecture. I didn’t have much involvement in things outside SQL Server but this caused many, many things to change in our environment.  Many applications today aren’t just executables or web sites.  They are a combination of those plus network infrastructure, reports, network ports, IP addresses, DTS and SSIS packages, batch systems and many other things.  These all needed a little bit of attention to make sure they were functioning properly. Profiler turned out to be a handy tool.  I started a trace for failed logins and kept that running.  That let me fix a number of problems before people were able to report them.  I also ran traces to capture exceptions.  This helped identify problems with linked servers. Overall the thing that gave me the most problem was linked servers.  In order for a linked server to function properly you need to be pointed to the right server, have the proper login information, have the network routes available and have MSDTC configured properly.  We have a lot of linked servers and this created many failure points.  Some of the older linked servers used IP addresses and not DNS names.  This meant we had to go in and touch all those linked servers when the servers moved.

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  • SQL Server service accounts and SPNs

    - by simonsabin
    Service Principal Names (SPNs) are a must for kerberos authentication which is a must when using sharepoint, reporting services and sql server where you access one server that then needs to access another resource, this is called the double hop. The reason this is a complex problem is that the second hop has to be done with impersonation/delegation. For this to work there needs to be a way for the security system to make sure that the service in the middle is allowed to impersonate you, after all you are not giving the service your password. To do this you need to be using kerberos. The following is my simple interpretation of how kerberos works. I find the Kerberos documentation rediculously complex so the following might be sligthly wrong but I think its close enough. Keberos works on a ticketing system, the prinicipal is that you get a security token from AD and then you can pass that to the service in the middle which can then use that token to impersonate you. For that to work AD has to be able to identify who is allowed to use the token, in this case the service account.But how do you as a client know what service account the service in the middle is configured with. The answer is SPNs. The SPN is the mapping between your logical connection to the service account. One type of SPN is for the DNS name for the server and the port. i.e. MySQL.mydomain.com and 1433. You can see how this maps to SQL Server on that server, but how does it map to the account. Well it can be done in two ways, either you can have a mapping defined in AD or AD can use a default mapping (this is something I didn't know about). To map the SPN in AD then you have to add the SPN to the user account, this is documented in the first link below either directly or using a tool called SetSPN. You might say that is complex, well it is and thats why SQL Server tries to do it for you, at start up it tries to connect to AD and set the SPN on the account it is running as, clearly that can only happen IF SQL is running as a domain account AND importantly it has permission to do so. By default a normal domain user account doesn't have the correct permission, and is why so many people have this problem. If the account is a domain admin then it will have permission, but non of us run SQL using domain admin accounts do we. You might also note that the SPN contains the port number (this isn't a requirement now in sql 2008 but I won't go into that), so if you set it manually and you are using dynamic ports (the default for a named instance) what do you do, well every time the port changes you need to change the SPN allocated to the account. Thats why its advised to let SQL Server register the SPN itself. You may also have thought, well what happens if I change my service account, won't that lead to two accounts with the same SPN. Possibly. Having two accounts with the same SPN is definitely a problem. Why? Well because if there are two accounts Kerberos can't identify the exact account that the service is running as, it could be either account, and so your security falls back to NTLM. SETSPN is useful for finding duplicate SPNs Reading this you will probably be thinking Oh my goodness this is really difficult. It is however I've found today in investigating something else that there is an easy option. Use Network Service as your service account. Network Service is a special account and is tied to the computer. It appears that Network Service has the update rights to AD to set an SPN mapping for the computer account. This then allows the SPN mapping to work. I believe this also works for the local system account. To get all the SPNs in your AD run the following, it could be a large file, so you might want to restrict it to a specific OU, or CN ldifde -d "DC=<domain>" -l servicePrincipalName -F spn.txt You will read in the links below that you need SQL to register the SPN this is done how to use Kerberos authenticaiton in SQL Server - http://support.microsoft.com/kb/319723 Using Kerberos with SQL Server - http://blogs.msdn.com/sql_protocols/archive/2005/10/12/479871.aspx Understanding Kerberos and NTLM authentication in SQL Server Connections - http://blogs.msdn.com/sql_protocols/archive/2006/12/02/understanding-kerberos-and-ntlm-authentication-in-sql-server-connections.aspx Summary The only reason I personally know to use a domain account is when you can't get kerberos to work and you want to do BULK INSERT or other network service that requires access to a a remote server. In this case you have to resort to using SQL authentication and the SQL Server uses its service account to access the remote service, and thus you need a domain account. You migth need this if using some forms of replication. I've always found Kerberos awkward to setup and so fallen back to this domain account approach. So in summary to get Kerberos to work try using the network service or local system accounts. For a great post from the Adam Saxton of the SQL Server support team go to http://blogs.msdn.com/psssql/archive/2010/03/09/what-spn-do-i-use-and-how-does-it-get-there.aspx 

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  • Sorting Algorithms

    - by MarkPearl
    General Every time I go back to university I find myself wading through sorting algorithms and their implementation in C++. Up to now I haven’t really appreciated their true value. However as I discovered this last week with Dictionaries in C# – having a knowledge of some basic programming principles can greatly improve the performance of a system and make one think twice about how to tackle a problem. I’m going to cover briefly in this post the following: Selection Sort Insertion Sort Shellsort Quicksort Mergesort Heapsort (not complete) Selection Sort Array based selection sort is a simple approach to sorting an unsorted array. Simply put, it repeats two basic steps to achieve a sorted collection. It starts with a collection of data and repeatedly parses it, each time sorting out one element and reducing the size of the next iteration of parsed data by one. So the first iteration would go something like this… Go through the entire array of data and find the lowest value Place the value at the front of the array The second iteration would go something like this… Go through the array from position two (position one has already been sorted with the smallest value) and find the next lowest value in the array. Place the value at the second position in the array This process would be completed until the entire array had been sorted. A positive about selection sort is that it does not make many item movements. In fact, in a worst case scenario every items is only moved once. Selection sort is however a comparison intensive sort. If you had 10 items in a collection, just to parse the collection you would have 10+9+8+7+6+5+4+3+2=54 comparisons to sort regardless of how sorted the collection was to start with. If you think about it, if you applied selection sort to a collection already sorted, you would still perform relatively the same number of iterations as if it was not sorted at all. Many of the following algorithms try and reduce the number of comparisons if the list is already sorted – leaving one with a best case and worst case scenario for comparisons. Likewise different approaches have different levels of item movement. Depending on what is more expensive, one may give priority to one approach compared to another based on what is more expensive, a comparison or a item move. Insertion Sort Insertion sort tries to reduce the number of key comparisons it performs compared to selection sort by not “doing anything” if things are sorted. Assume you had an collection of numbers in the following order… 10 18 25 30 23 17 45 35 There are 8 elements in the list. If we were to start at the front of the list – 10 18 25 & 30 are already sorted. Element 5 (23) however is smaller than element 4 (30) and so needs to be repositioned. We do this by copying the value at element 5 to a temporary holder, and then begin shifting the elements before it up one. So… Element 5 would be copied to a temporary holder 10 18 25 30 23 17 45 35 – T 23 Element 4 would shift to Element 5 10 18 25 30 30 17 45 35 – T 23 Element 3 would shift to Element 4 10 18 25 25 30 17 45 35 – T 23 Element 2 (18) is smaller than the temporary holder so we put the temporary holder value into Element 3. 10 18 23 25 30 17 45 35 – T 23   We now have a sorted list up to element 6. And so we would repeat the same process by moving element 6 to a temporary value and then shifting everything up by one from element 2 to element 5. As you can see, one major setback for this technique is the shifting values up one – this is because up to now we have been considering the collection to be an array. If however the collection was a linked list, we would not need to shift values up, but merely remove the link from the unsorted value and “reinsert” it in a sorted position. Which would reduce the number of transactions performed on the collection. So.. Insertion sort seems to perform better than selection sort – however an implementation is slightly more complicated. This is typical with most sorting algorithms – generally, greater performance leads to greater complexity. Also, insertion sort performs better if a collection of data is already sorted. If for instance you were handed a sorted collection of size n, then only n number of comparisons would need to be performed to verify that it is sorted. It’s important to note that insertion sort (array based) performs a number item moves – every time an item is “out of place” several items before it get shifted up. Shellsort – Diminishing Increment Sort So up to now we have covered Selection Sort & Insertion Sort. Selection Sort makes many comparisons and insertion sort (with an array) has the potential of making many item movements. Shellsort is an approach that takes the normal insertion sort and tries to reduce the number of item movements. In Shellsort, elements in a collection are viewed as sub-collections of a particular size. Each sub-collection is sorted so that the elements that are far apart move closer to their final position. Suppose we had a collection of 15 elements… 10 20 15 45 36 48 7 60 18 50 2 19 43 30 55 First we may view the collection as 7 sub-collections and sort each sublist, lets say at intervals of 7 10 60 55 – 20 18 – 15 50 – 45 2 – 36 19 – 48 43 – 7 30 10 55 60 – 18 20 – 15 50 – 2 45 – 19 36 – 43 48 – 7 30 (Sorted) We then sort each sublist at a smaller inter – lets say 4 10 55 60 18 – 20 15 50 2 – 45 19 36 43 – 48 7 30 10 18 55 60 – 2 15 20 50 – 19 36 43 45 – 7 30 48 (Sorted) We then sort elements at a distance of 1 (i.e. we apply a normal insertion sort) 10 18 55 60 2 15 20 50 19 36 43 45 7 30 48 2 7 10 15 18 19 20 30 36 43 45 48 50 55 (Sorted) The important thing with shellsort is deciding on the increment sequence of each sub-collection. From what I can tell, there isn’t any definitive method and depending on the order of your elements, different increment sequences may perform better than others. There are however certain increment sequences that you may want to avoid. An even based increment sequence (e.g. 2 4 8 16 32 …) should typically be avoided because it does not allow for even elements to be compared with odd elements until the final sort phase – which in a way would negate many of the benefits of using sub-collections. The performance on the number of comparisons and item movements of Shellsort is hard to determine, however it is considered to be considerably better than the normal insertion sort. Quicksort Quicksort uses a divide and conquer approach to sort a collection of items. The collection is divided into two sub-collections – and the two sub-collections are sorted and combined into one list in such a way that the combined list is sorted. The algorithm is in general pseudo code below… Divide the collection into two sub-collections Quicksort the lower sub-collection Quicksort the upper sub-collection Combine the lower & upper sub-collection together As hinted at above, quicksort uses recursion in its implementation. The real trick with quicksort is to get the lower and upper sub-collections to be of equal size. The size of a sub-collection is determined by what value the pivot is. Once a pivot is determined, one would partition to sub-collections and then repeat the process on each sub collection until you reach the base case. With quicksort, the work is done when dividing the sub-collections into lower & upper collections. The actual combining of the lower & upper sub-collections at the end is relatively simple since every element in the lower sub-collection is smaller than the smallest element in the upper sub-collection. Mergesort With quicksort, the average-case complexity was O(nlog2n) however the worst case complexity was still O(N*N). Mergesort improves on quicksort by always having a complexity of O(nlog2n) regardless of the best or worst case. So how does it do this? Mergesort makes use of the divide and conquer approach to partition a collection into two sub-collections. It then sorts each sub-collection and combines the sorted sub-collections into one sorted collection. The general algorithm for mergesort is as follows… Divide the collection into two sub-collections Mergesort the first sub-collection Mergesort the second sub-collection Merge the first sub-collection and the second sub-collection As you can see.. it still pretty much looks like quicksort – so lets see where it differs… Firstly, mergesort differs from quicksort in how it partitions the sub-collections. Instead of having a pivot – merge sort partitions each sub-collection based on size so that the first and second sub-collection of relatively the same size. This dividing keeps getting repeated until the sub-collections are the size of a single element. If a sub-collection is one element in size – it is now sorted! So the trick is how do we put all these sub-collections together so that they maintain their sorted order. Sorted sub-collections are merged into a sorted collection by comparing the elements of the sub-collection and then adjusting the sorted collection. Lets have a look at a few examples… Assume 2 sub-collections with 1 element each 10 & 20 Compare the first element of the first sub-collection with the first element of the second sub-collection. Take the smallest of the two and place it as the first element in the sorted collection. In this scenario 10 is smaller than 20 so 10 is taken from sub-collection 1 leaving that sub-collection empty, which means by default the next smallest element is in sub-collection 2 (20). So the sorted collection would be 10 20 Lets assume 2 sub-collections with 2 elements each 10 20 & 15 19 So… again we would Compare 10 with 15 – 10 is the winner so we add it to our sorted collection (10) leaving us with 20 & 15 19 Compare 20 with 15 – 15 is the winner so we add it to our sorted collection (10 15) leaving us with 20 & 19 Compare 20 with 19 – 19 is the winner so we add it to our sorted collection (10 15 19) leaving us with 20 & _ 20 is by default the winner so our sorted collection is 10 15 19 20. Make sense? Heapsort (still needs to be completed) So by now I am tired of sorting algorithms and trying to remember why they were so important. I think every year I go through this stuff I wonder to myself why are we made to learn about selection sort and insertion sort if they are so bad – why didn’t we just skip to Mergesort & Quicksort. I guess the only explanation I have for this is that sometimes you learn things so that you can implement them in future – and other times you learn things so that you know it isn’t the best way of implementing things and that you don’t need to implement it in future. Anyhow… luckily this is going to be the last one of my sorts for today. The first step in heapsort is to convert a collection of data into a heap. After the data is converted into a heap, sorting begins… So what is the definition of a heap? If we have to convert a collection of data into a heap, how do we know when it is a heap and when it is not? The definition of a heap is as follows: A heap is a list in which each element contains a key, such that the key in the element at position k in the list is at least as large as the key in the element at position 2k +1 (if it exists) and 2k + 2 (if it exists). Does that make sense? At first glance I’m thinking what the heck??? But then after re-reading my notes I see that we are doing something different – up to now we have really looked at data as an array or sequential collection of data that we need to sort – a heap represents data in a slightly different way – although the data is stored in a sequential collection, for a sequential collection of data to be in a valid heap – it is “semi sorted”. Let me try and explain a bit further with an example… Example 1 of Potential Heap Data Assume we had a collection of numbers as follows 1[1] 2[2] 3[3] 4[4] 5[5] 6[6] For this to be a valid heap element with value of 1 at position [1] needs to be greater or equal to the element at position [3] (2k +1) and position [4] (2k +2). So in the above example, the collection of numbers is not in a valid heap. Example 2 of Potential Heap Data Lets look at another collection of numbers as follows 6[1] 5[2] 4[3] 3[4] 2[5] 1[6] Is this a valid heap? Well… element with the value 6 at position 1 must be greater or equal to the element at position [3] and position [4]. Is 6 > 4 and 6 > 3? Yes it is. Lets look at element 5 as position 2. It must be greater than the values at [4] & [5]. Is 5 > 3 and 5 > 2? Yes it is. If you continued to examine this second collection of data you would find that it is in a valid heap based on the definition of a heap.

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  • My History with Agile

    - by Robert May
    I’m going to write my history with Agile here.  That way, in future posts, I can refer back to it, instead of typing it out in the post that contains information you may actually want to read.  Note that I’m actually a pretty senior developer, and do lots of technical interviews.  I’m an Agile fan because of the difference it makes in peoples lives and the improvement in quality it brings, and I’ll sacrifice my technological advance to help teams. Management History I started management pretty early in my career, starting with the first job that I ever had.  I actually do NOT have a CS or similar degree.  I have a Bachelor’s of Business Administration with an emphasis in Computer Information Systems. My first management gigs were around call center work and were very schedule oriented.  I didn’t understand the true value of teams, and I’m ashamed to admit, I actually installed a fingerprint scanner as a time clock in this job.  I shudder to think of the impact that I had on the team spirit.  I didn’t even trust them enough to fill out their time cards correctly.  How sad. I was managing nearly 100 people in this position, with the help of a great set of subordinates. I did try to come up with reward programs for the team, but again, didn’t understand the concept of team, so instead of letting the team determine how the rewards should work, I mandated from on high, which isn’t a good thing. I was told that I wasn’t the type that would be a good manager by people whom I respected a lot.  They said it because I was a computer geek, since they don’t understand good management either, but in retrospect, they were right about me then.  I was too green. After my first job, I went on to other jobs and with the exception of one job, I’ve managed people at them all.  The rest of the management story is important for understanding agile, so I’ll save it for my next post. Technical History I’ve been in software development for many, many years.  I technically started programming on a commodore 64 in basic.  I didn’t know that I was programming, but I was sure having fun.  That was followed by batch files, Gorilla hacking (I always had to win), WordPerfect Macro programming and other things that taught me the basics. My first “real” job was with a telephone company, and that’s where I made my first database application in DataEase, wrote my first VBA app and started using real programming tools, like turbo pascal, vb3-vb5, and semi-real tools like RPG and VisualRPG.  I wrote my first web page in 1994, and built my first data driven web page in 1995 using perlDB.  You really can do anything with Perl.  At this time, I also started a Linux based internet service provider that is still in operation today.  One of the people I worked with is now a Microsoft employee building and designing frameworks you probably know well.  Smart guy.  I also built my first ASP applications connecting to Sql Server 6.5, setup Exchange 5.5 for the company, and many other system administration stuff.  I’m a programmer by choice, mostly because I don’t really like PC support. From there, I went on to a large state agency.  I got to see and maintain true waterfall projects.  5 years of maintaining the 200 VB COM+ (MTS, actually) dlls that were used to calculate a single number is a long time.  That was all Microsoft DNS technologies.  SQL Server and VB6 were the tools of choice, although .net started to be a factor near the end of employment.  I did some heavy XML work at this job and even wrote an XSD parser and validator in VB6 that was a shim until MSXML 3.0 came out.  Prior to 3.0, XSD’s weren’t supported, and I didn’t want to write DTDs. Ironically, jobs after this were more generic.  I pretty much settled in on the .net framework and revisions of it.  Lots of WPF, some silverlight, lots of ASP.NET, some SQL Azure, lots of SQL Server, some Oracle, but I don’t think that I was as passionate about development and technologies.  I was more into the management of development.  I like people. Technorati Tags: Agile,history

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  • Plan Caching and Query Memory Part II (Hash Match) – When not to use stored procedure - Most common performance mistake SQL Server developers make.

    - by sqlworkshops
    SQL Server estimates Memory requirement at compile time, when stored procedure or other plan caching mechanisms like sp_executesql or prepared statement are used, the memory requirement is estimated based on first set of execution parameters. This is a common reason for spill over tempdb and hence poor performance. Common memory allocating queries are that perform Sort and do Hash Match operations like Hash Join or Hash Aggregation or Hash Union. This article covers Hash Match operations with examples. It is recommended to read Plan Caching and Query Memory Part I before this article which covers an introduction and Query memory for Sort. In most cases it is cheaper to pay for the compilation cost of dynamic queries than huge cost for spill over tempdb, unless memory requirement for a query does not change significantly based on predicates.   This article covers underestimation / overestimation of memory for Hash Match operation. Plan Caching and Query Memory Part I covers underestimation / overestimation for Sort. It is important to note that underestimation of memory for Sort and Hash Match operations lead to spill over tempdb and hence negatively impact performance. Overestimation of memory affects the memory needs of other concurrently executing queries. In addition, it is important to note, with Hash Match operations, overestimation of memory can actually lead to poor performance.   To read additional articles I wrote click here.   The best way to learn is to practice. To create the below tables and reproduce the behavior, join the mailing list by using this link: www.sqlworkshops.com/ml and I will send you the table creation script. Most of these concepts are also covered in our webcasts: www.sqlworkshops.com/webcasts  Let’s create a Customer’s State table that has 99% of customers in NY and the rest 1% in WA.Customers table used in Part I of this article is also used here.To observe Hash Warning, enable 'Hash Warning' in SQL Profiler under Events 'Errors and Warnings'. --Example provided by www.sqlworkshops.com drop table CustomersState go create table CustomersState (CustomerID int primary key, Address char(200), State char(2)) go insert into CustomersState (CustomerID, Address) select CustomerID, 'Address' from Customers update CustomersState set State = 'NY' where CustomerID % 100 != 1 update CustomersState set State = 'WA' where CustomerID % 100 = 1 go update statistics CustomersState with fullscan go   Let’s create a stored procedure that joins customers with CustomersState table with a predicate on State. --Example provided by www.sqlworkshops.com create proc CustomersByState @State char(2) as begin declare @CustomerID int select @CustomerID = e.CustomerID from Customers e inner join CustomersState es on (e.CustomerID = es.CustomerID) where es.State = @State option (maxdop 1) end go  Let’s execute the stored procedure first with parameter value ‘WA’ – which will select 1% of data. set statistics time on go --Example provided by www.sqlworkshops.com exec CustomersByState 'WA' goThe stored procedure took 294 ms to complete.  The stored procedure was granted 6704 KB based on 8000 rows being estimated.  The estimated number of rows, 8000 is similar to actual number of rows 8000 and hence the memory estimation should be ok.  There was no Hash Warning in SQL Profiler. To observe Hash Warning, enable 'Hash Warning' in SQL Profiler under Events 'Errors and Warnings'.   Now let’s execute the stored procedure with parameter value ‘NY’ – which will select 99% of data. -Example provided by www.sqlworkshops.com exec CustomersByState 'NY' go  The stored procedure took 2922 ms to complete.   The stored procedure was granted 6704 KB based on 8000 rows being estimated.    The estimated number of rows, 8000 is way different from the actual number of rows 792000 because the estimation is based on the first set of parameter value supplied to the stored procedure which is ‘WA’ in our case. This underestimation will lead to spill over tempdb, resulting in poor performance.   There was Hash Warning (Recursion) in SQL Profiler. To observe Hash Warning, enable 'Hash Warning' in SQL Profiler under Events 'Errors and Warnings'.   Let’s recompile the stored procedure and then let’s first execute the stored procedure with parameter value ‘NY’.  In a production instance it is not advisable to use sp_recompile instead one should use DBCC FREEPROCCACHE (plan_handle). This is due to locking issues involved with sp_recompile, refer to our webcasts, www.sqlworkshops.com/webcasts for further details.   exec sp_recompile CustomersByState go --Example provided by www.sqlworkshops.com exec CustomersByState 'NY' go  Now the stored procedure took only 1046 ms instead of 2922 ms.   The stored procedure was granted 146752 KB of memory. The estimated number of rows, 792000 is similar to actual number of rows of 792000. Better performance of this stored procedure execution is due to better estimation of memory and avoiding spill over tempdb.   There was no Hash Warning in SQL Profiler.   Now let’s execute the stored procedure with parameter value ‘WA’. --Example provided by www.sqlworkshops.com exec CustomersByState 'WA' go  The stored procedure took 351 ms to complete, higher than the previous execution time of 294 ms.    This stored procedure was granted more memory (146752 KB) than necessary (6704 KB) based on parameter value ‘NY’ for estimation (792000 rows) instead of parameter value ‘WA’ for estimation (8000 rows). This is because the estimation is based on the first set of parameter value supplied to the stored procedure which is ‘NY’ in this case. This overestimation leads to poor performance of this Hash Match operation, it might also affect the performance of other concurrently executing queries requiring memory and hence overestimation is not recommended.     The estimated number of rows, 792000 is much more than the actual number of rows of 8000.  Intermediate Summary: This issue can be avoided by not caching the plan for memory allocating queries. Other possibility is to use recompile hint or optimize for hint to allocate memory for predefined data range.Let’s recreate the stored procedure with recompile hint. --Example provided by www.sqlworkshops.com drop proc CustomersByState go create proc CustomersByState @State char(2) as begin declare @CustomerID int select @CustomerID = e.CustomerID from Customers e inner join CustomersState es on (e.CustomerID = es.CustomerID) where es.State = @State option (maxdop 1, recompile) end go  Let’s execute the stored procedure initially with parameter value ‘WA’ and then with parameter value ‘NY’. --Example provided by www.sqlworkshops.com exec CustomersByState 'WA' go exec CustomersByState 'NY' go  The stored procedure took 297 ms and 1102 ms in line with previous optimal execution times.   The stored procedure with parameter value ‘WA’ has good estimation like before.   Estimated number of rows of 8000 is similar to actual number of rows of 8000.   The stored procedure with parameter value ‘NY’ also has good estimation and memory grant like before because the stored procedure was recompiled with current set of parameter values.  Estimated number of rows of 792000 is similar to actual number of rows of 792000.    The compilation time and compilation CPU of 1 ms is not expensive in this case compared to the performance benefit.   There was no Hash Warning in SQL Profiler.   Let’s recreate the stored procedure with optimize for hint of ‘NY’. --Example provided by www.sqlworkshops.com drop proc CustomersByState go create proc CustomersByState @State char(2) as begin declare @CustomerID int select @CustomerID = e.CustomerID from Customers e inner join CustomersState es on (e.CustomerID = es.CustomerID) where es.State = @State option (maxdop 1, optimize for (@State = 'NY')) end go  Let’s execute the stored procedure initially with parameter value ‘WA’ and then with parameter value ‘NY’. --Example provided by www.sqlworkshops.com exec CustomersByState 'WA' go exec CustomersByState 'NY' go  The stored procedure took 353 ms with parameter value ‘WA’, this is much slower than the optimal execution time of 294 ms we observed previously. This is because of overestimation of memory. The stored procedure with parameter value ‘NY’ has optimal execution time like before.   The stored procedure with parameter value ‘WA’ has overestimation of rows because of optimize for hint value of ‘NY’.   Unlike before, more memory was estimated to this stored procedure based on optimize for hint value ‘NY’.    The stored procedure with parameter value ‘NY’ has good estimation because of optimize for hint value of ‘NY’. Estimated number of rows of 792000 is similar to actual number of rows of 792000.   Optimal amount memory was estimated to this stored procedure based on optimize for hint value ‘NY’.   There was no Hash Warning in SQL Profiler.   This article covers underestimation / overestimation of memory for Hash Match operation. Plan Caching and Query Memory Part I covers underestimation / overestimation for Sort. It is important to note that underestimation of memory for Sort and Hash Match operations lead to spill over tempdb and hence negatively impact performance. Overestimation of memory affects the memory needs of other concurrently executing queries. In addition, it is important to note, with Hash Match operations, overestimation of memory can actually lead to poor performance.   Summary: Cached plan might lead to underestimation or overestimation of memory because the memory is estimated based on first set of execution parameters. It is recommended not to cache the plan if the amount of memory required to execute the stored procedure has a wide range of possibilities. One can mitigate this by using recompile hint, but that will lead to compilation overhead. However, in most cases it might be ok to pay for compilation rather than spilling sort over tempdb which could be very expensive compared to compilation cost. The other possibility is to use optimize for hint, but in case one sorts more data than hinted by optimize for hint, this will still lead to spill. On the other side there is also the possibility of overestimation leading to unnecessary memory issues for other concurrently executing queries. In case of Hash Match operations, this overestimation of memory might lead to poor performance. When the values used in optimize for hint are archived from the database, the estimation will be wrong leading to worst performance, so one has to exercise caution before using optimize for hint, recompile hint is better in this case.   I explain these concepts with detailed examples in my webcasts (www.sqlworkshops.com/webcasts), I recommend you to watch them. The best way to learn is to practice. To create the above tables and reproduce the behavior, join the mailing list at www.sqlworkshops.com/ml and I will send you the relevant SQL Scripts.  Register for the upcoming 3 Day Level 400 Microsoft SQL Server 2008 and SQL Server 2005 Performance Monitoring & Tuning Hands-on Workshop in London, United Kingdom during March 15-17, 2011, click here to register / Microsoft UK TechNet.These are hands-on workshops with a maximum of 12 participants and not lectures. For consulting engagements click here.   Disclaimer and copyright information:This article refers to organizations and products that may be the trademarks or registered trademarks of their various owners. Copyright of this article belongs to R Meyyappan / www.sqlworkshops.com. You may freely use the ideas and concepts discussed in this article with acknowledgement (www.sqlworkshops.com), but you may not claim any of it as your own work. This article is for informational purposes only; you use any of the suggestions given here entirely at your own risk.   R Meyyappan [email protected] LinkedIn: http://at.linkedin.com/in/rmeyyappan

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  • .NET Security Part 2

    - by Simon Cooper
    So, how do you create partial-trust appdomains? Where do you come across them? There are two main situations in which your assembly runs as partially-trusted using the Microsoft .NET stack: Creating a CLR assembly in SQL Server with anything other than the UNSAFE permission set. The permissions available in each permission set are given here. Loading an assembly in ASP.NET in any trust level other than Full. Information on ASP.NET trust levels can be found here. You can configure the specific permissions available to assemblies using ASP.NET policy files. Alternatively, you can create your own partially-trusted appdomain in code and directly control the permissions and the full-trust API available to the assemblies you load into the appdomain. This is the scenario I’ll be concentrating on in this post. Creating a partially-trusted appdomain There is a single overload of AppDomain.CreateDomain that allows you to specify the permissions granted to assemblies in that appdomain – this one. This is the only call that allows you to specify a PermissionSet for the domain. All the other calls simply use the permissions of the calling code. If the permissions are restricted, then the resulting appdomain is referred to as a sandboxed domain. There are three things you need to create a sandboxed domain: The specific permissions granted to all assemblies in the domain. The application base (aka working directory) of the domain. The list of assemblies that have full-trust if they are loaded into the sandboxed domain. The third item is what allows us to have a fully-trusted API that is callable by partially-trusted code. I’ll be looking at the details of this in a later post. Granting permissions to the appdomain Firstly, the permissions granted to the appdomain. This is encapsulated in a PermissionSet object, initialized either with no permissions or full-trust permissions. For sandboxed appdomains, the PermissionSet is initialized with no permissions, then you add permissions you want assemblies loaded into that appdomain to have by default: PermissionSet restrictedPerms = new PermissionSet(PermissionState.None); // all assemblies need Execution permission to run at all restrictedPerms.AddPermission( new SecurityPermission(SecurityPermissionFlag.Execution)); // grant general read access to C:\config.xml restrictedPerms.AddPermission( new FileIOPermission(FileIOPermissionAccess.Read, @"C:\config.xml")); // grant permission to perform DNS lookups restrictedPerms.AddPermission( new DnsPermission(PermissionState.Unrestricted)); It’s important to point out that the permissions granted to an appdomain, and so to all assemblies loaded into that appdomain, are usable without needing to go through any SafeCritical code (see my last post if you’re unsure what SafeCritical code is). That is, partially-trusted code loaded into an appdomain with the above permissions (and so running under the Transparent security level) is able to create and manipulate a FileStream object to read from C:\config.xml directly. It is only for operations requiring permissions that are not granted to the appdomain that partially-trusted code is required to call a SafeCritical method that then asserts the missing permissions and performs the operation safely on behalf of the partially-trusted code. The application base of the domain This is simply set as a property on an AppDomainSetup object, and is used as the default directory assemblies are loaded from: AppDomainSetup appDomainSetup = new AppDomainSetup { ApplicationBase = @"C:\temp\sandbox", }; If you’ve read the documentation around sandboxed appdomains, you’ll notice that it mentions a security hole if this parameter is set correctly. I’ll be looking at this, and other pitfalls, that will break the sandbox when using sandboxed appdomains, in a later post. Full-trust assemblies in the appdomain Finally, we need the strong names of the assemblies that, when loaded into the appdomain, will be run as full-trust, irregardless of the permissions specified on the appdomain. These assemblies will contain methods and classes decorated with SafeCritical and Critical attributes. I’ll be covering the details of creating full-trust APIs for partial-trust appdomains in a later post. This is how you get the strongnames of an assembly to be executed as full-trust in the sandbox: // get the Assembly object for the assembly Assembly assemblyWithApi = ... // get the StrongName from the assembly's collection of evidence StrongName apiStrongName = assemblyWithApi.Evidence.GetHostEvidence<StrongName>(); Creating the sandboxed appdomain So, putting these three together, you create the appdomain like so: AppDomain sandbox = AppDomain.CreateDomain( "Sandbox", null, appDomainSetup, restrictedPerms, apiStrongName); You can then load and execute assemblies in this appdomain like any other. For example, to load an assembly into the appdomain and get an instance of the Sandboxed.Entrypoint class, implementing IEntrypoint, you do this: IEntrypoint o = (IEntrypoint)sandbox.CreateInstanceFromAndUnwrap( "C:\temp\sandbox\SandboxedAssembly.dll", "Sandboxed.Entrypoint"); // call method the Execute method on this object within the sandbox o.Execute(); The second parameter to CreateDomain is for security evidence used in the appdomain. This was a feature of the .NET 2 security model, and has been (mostly) obsoleted in the .NET 4 model. Unless the evidence is needed elsewhere (eg. isolated storage), you can pass in null for this parameter. Conclusion That’s the basics of sandboxed appdomains. The most important object is the PermissionSet that defines the permissions available to assemblies running in the appdomain; it is this object that defines the appdomain as full or partial-trust. The appdomain also needs a default directory used for assembly lookups as the ApplicationBase parameter, and you can specify an optional list of the strongnames of assemblies that will be given full-trust permissions if they are loaded into the sandboxed appdomain. Next time, I’ll be looking closer at full-trust assemblies running in a sandboxed appdomain, and what you need to do to make an API available to partial-trust code.

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  • Configure IPv6 on your Linux system (Ubuntu)

    After the presentation on IPv6 at the first event of the Emtel Knowledge Series and some recent discussion on social media networks with other geeks and Linux interested IT people here in Mauritius, I thought that I should give it a try (finally) and tweak my local network infrastructure. Honestly, I have been to busy with contractual project work and it never really occurred to me to set up IPv6 in my LAN. Well, the following paragraphs are going to shed some light on those aspects of modern computer and network technology. This is the first article in a series on IPv6 configuration: Configure IPv6 on your Linux system DHCPv6: Provide IPv6 information in your local network Enabling DNS for IPv6 infrastructure Accessing your web server via IPv6 Piece of advice: This is based on my findings on the internet while reading other people's helpful articles and going through a couple of man-pages on my local system. Let's embrace IPv6 The basic configuration on Linux is actually very simple as the kernel, operating system, and user-space programs support that protocol natively. If your system is ready to go for IP (aka: IPv4), then you are good to go for anything else. At least, I didn't have to install any additional packages on my system(s). We are going to assign a static IPv6 address to the system. Hence, we have to modify the definition of interfaces and check whether we have an inet6 entry specified. Open your favourite text editor and check the following entries (it should be at least similar to this): $ sudo nano /etc/network/interfaces auto eth0# IPv4 configurationiface eth0 inet static  address 192.168.1.2  network 192.168.1.0  netmask 255.255.255.0  broadcast 192.168.1.255# IPv6 configurationiface eth0 inet6 static  pre-up modprobe ipv6  address 2001:db8:bad:a55::2  netmask 64 Of course, you might have to adjust your interface device (eth0) or you might be interested to have multiple directives for additional devices (eth1, eth2, etc.). The auto instruction takes care that your device is enabled and configured during the booting phase. The use of the pre-up directive depends on your kernel configuration but in most scenarios this might be an optional line. Anyways, it doesn't hurt to have it enabled after all - just to be on the safe side. Next, either restart your network subsystem like so: $ sudo service networking restart Or you might prefer to do it manually with identical parameters, like so: $ sudo ifconfig eth0 inet6 add 2001:db8:bad:a55::2/64 In case that you're logged in remotely into your PC (ie. via ssh), it is highly advised to opt for the second choice and add the device manually. You can check your configuration afterwards with one of the following commands (depends on whether it is installed): $ sudo ifconfig eth0eth0      Link encap:Ethernet  HWaddr 00:21:5a:50:d7:94            inet addr:192.168.160.2  Bcast:192.168.160.255  Mask:255.255.255.0          inet6 addr: fe80::221:5aff:fe50:d794/64 Scope:Link          inet6 addr: 2001:db8:bad:a55::2/64 Scope:Global          UP BROADCAST RUNNING MULTICAST  MTU:1500  Metric:1 $ sudo ip -6 address show eth03: eth0: <BROADCAST,MULTICAST,UP,LOWER_UP> mtu 1500 qlen 1000    inet6 2001:db8:bad:a55::2/64 scope global        valid_lft forever preferred_lft forever    inet6 fe80::221:5aff:fe50:d794/64 scope link        valid_lft forever preferred_lft forever In both cases, it confirms that our network device has been assigned a valid IPv6 address. That's it in general for your setup on one system. But of course, you might be interested to enable more services for IPv6, especially if you're already running a couple of them in your IP network. More details are available on the official Ubuntu Wiki. Continue to configure your network to provide IPv6 address information automatically in your local infrastructure.

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  • Easy Made Easier

    - by dragonfly
        How easy is it to deploy a 2 node, fully redundant Oracle RAC cluster? Not very. Unless you use an Oracle Database Appliance. The focus of this member of Oracle's Engineered Systems family is to simplify the configuration, management and maintenance throughout the life of the system, while offering pay-as-you-grow scaling. Getting a 2-node RAC cluster up and running in under 2 hours has been made possible by the Oracle Database Appliance. Don't take my word for it, just check out these blog posts from partners and end users. The Oracle Database Appliance Experience - Zip Zoom Zoom http://www.fuadarshad.com/2012/02/oracle-database-appliance-experience.html Off-the-shelf Oracle database servers http://normanweaver.wordpress.com/2011/10/10/off-the-shelf-oracle-database-servers/ Oracle Database Appliance – Deployment Steps http://marcel.vandewaters.nl/oracle/database-appliance/oracle-database-appliance-deployment-steps     See how easy it is to deploy an Oracle Database Appliance for high availability with RAC? Now for the meat of this post, which is the first in a series of posts describing tips for making the deployment of an ODA even easier. The key to the easy deployment of an Oracle Database Appliance is the Appliance Manager software, which does the actual software deployment and configuration, based on best practices. But in order for it to do that, it needs some basic information first, including system name, IP addresses, etc. That's where the Appliance Manager GUI comes in to play, taking a wizard approach to specifying the information needed.     Using the Appliance Manager GUI is pretty straight forward, stepping through several screens of information to enter data in typical wizard style. Like most configuration tasks, it helps to gather the required information before hand. But before you rush out to a committee meeting on what to use for host names, and rely on whatever IP addresses might be hanging around, make sure you are familiar with some of the auto-fill defaults for the Appliance Manager. I'll step through the key screens below to highlight the results of the auto-fill capability of the Appliance Manager GUI.     Depending on which of the 2 Configuration Types (Config Type screen) you choose, you will get a slightly different set of screens. The Typical configuration assumes certain default configuration choices and has the fewest screens, where as the Custom configuration gives you the most flexibility in what you configure from the start. In the examples below, I have used the Custom config type.     One of the first items you are asked for is the System Name (System Info screen). This is used to identify the system, but also as the base for the default hostnames on following screens. In this screen shot, the System Name is "oda".     When you get to the next screen (Generic Network screen), you enter your domain name, DNS IP address(es), and NTP IP address(es). Next up is the Public Network screen, seen below, where you will see the host name fields are automatically filled in with default host names based on the System Name, in this case "oda". The System Name is also the basis for default host names for the extra ethernet ports available for configuration as part of a Custom configuration, as seen in the 2nd screen shot below (Other Network). There is no requirement to use these host names, as you can easily edit any of the host names. This does make filling in the configuration details easier and less prone to "fat fingers" if you are OK with these host names. Here is a full list of the automatically filled in host names. 1 2 1-vip 2-vip -scan 1-ilom 2-ilom 1-net1 2-net1 1-net2 2-net2 1-net3 2-net3     Another auto-fill feature of the Appliance Manager GUI follows a common practice of deploying IP Addresses for a RAC cluster in sequential order. In the screen shot below, I entered the first IP address (Node1-IP), then hit Tab to move to the next field. As a result, the next 5 IP address fields were automatically filled in with the next 5 IP addresses sequentially from the first one I entered. As with the host names, these are not required, and can be changed to whatever your IP address values are. One note of caution though, if the first IP Address field (Node1-IP) is filled out and you click in that field and back out, the following 5 IP addresses will be set to the sequential default. If you don't use the sequential IP addresses, pay attention to where you click that mouse. :-)     In the screen shot below, by entering the netmask value in the Netmask field, in this case 255.255.255.0, the gateway value was auto-filled into the Gateway field, based on the IP addresses and netmask previously entered. As always, you can change this value.     My last 2 screen shots illustrate that the same sequential IP address autofill and netmask to gateway autofill works when entering the IP configuration details for the Integrated Lights Out Manager (ILOM) for both nodes. The time these auto-fill capabilities save in entering data is nice, but from my perspective not as important as the opportunity to avoid data entry errors. In my next post in this series, I will touch on the benefit of using the network validation capability of the Appliance Manager GUI prior to deploying an Oracle Database Appliance.

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  • High Availability for IaaS, PaaS and SaaS in the Cloud

    - by BuckWoody
    Outages, natural disasters and unforeseen events have proved that even in a distributed architecture, you need to plan for High Availability (HA). In this entry I'll explain a few considerations for HA within Infrastructure-as-a-Service (IaaS), Platform-as-a-Service (PaaS) and Software-as-a-Service (SaaS). In a separate post I'll talk more about Disaster Recovery (DR), since each paradigm has a different way to handle that. Planning for HA in IaaS IaaS involves Virtual Machines - so in effect, an HA strategy here takes on many of the same characteristics as it would on-premises. The primary difference is that the vendor controls the hardware, so you need to verify what they do for things like local redundancy and so on from the hardware perspective. As far as what you can control and plan for, the primary factors fall into three areas: multiple instances, geographical dispersion and task-switching. In almost every cloud vendor I've studied, to ensure your application will be protected by any level of HA, you need to have at least two of the Instances (VM's) running. This makes sense, but you might assume that the vendor just takes care of that for you - they don't. If a single VM goes down (for whatever reason) then the access to it is lost. Depending on multiple factors, you might be able to recover the data, but you should assume that you can't. You should keep a sync to another location (perhaps the vendor's storage system in another geographic datacenter or to a local location) to ensure you can continue to serve your clients. You'll also need to host the same VM's in another geographical location. Everything from a vendor outage to a network path problem could prevent your users from reaching the system, so you need to have multiple locations to handle this. This means that you'll have to figure out how to manage state between the geo's. If the system goes down in the middle of a transaction, you need to figure out what part of the process the system was in, and then re-create or transfer that state to the second set of systems. If you didn't write the software yourself, this is non-trivial. You'll also need a manual or automatic process to detect the failure and re-route the traffic to your secondary location. You could flip a DNS entry (if your application can tolerate that) or invoke another process to alias the first system to the second, such as load-balancing and so on. There are many options, but all of them involve coding the state into the application layer. If you've simply moved a state-ful application to VM's, you may not be able to easily implement an HA solution. Planning for HA in PaaS Implementing HA in PaaS is a bit simpler, since it's built on the concept of stateless applications deployment. Once again, you need at least two copies of each element in the solution (web roles, worker roles, etc.) to remain available in a single datacenter. Also, you need to deploy the application again in a separate geo, but the advantage here is that you could work out a "shared storage" model such that state is auto-balanced across the world. In fact, you don't have to maintain a "DR" site, the alternate location can be live and serving clients, and only take on extra load if the other site is not available. In Windows Azure, you can use the Traffic Manager service top route the requests as a type of auto balancer. Even with these benefits, I recommend a second backup of storage in another geographic location. Storage is inexpensive; and that second copy can be used for not only HA but DR. Planning for HA in SaaS In Software-as-a-Service (such as Office 365, or Hadoop in Windows Azure) You have far less control over the HA solution, although you still maintain the responsibility to ensure you have it. Since each SaaS is different, check with the vendor on the solution for HA - and make sure you understand what they do and what you are responsible for. They may have no HA for that solution, or pin it to a particular geo, or perhaps they have a massive HA built in with automatic load balancing (which is often the case).   All of these options (with the exception of SaaS) involve higher costs for the design. Do not sacrifice reliability for cost - that will always cost you more in the end. Build in the redundancy and HA at the very outset of the project - if you try to tack it on later in the process the business will push back and potentially not implement HA. References: http://www.bing.com/search?q=windows+azure+High+Availability  (each type of implementation is different, so I'm routing you to a search on the topic - look for the "Patterns and Practices" results for the area in Azure you're interested in)

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  • Trouble with Samba Domain

    - by Arkevius
    I'm having a bit of trouble setting up this Samba domain correctly. I'm getting an Access Denied error when trying to add a Windows XP machine to the domain. I'll go through my scenario in detail, but for those of you wanting a TLDR summary it'll be at the bottom of this post. I have HP Proliant server with Ubuntu 12.04 LTS installed. For this particular environment, I need this server to act as a PDC, file server, and print server. I began by updating and upgrading the packages (of course). Then went to install samba, gnome-desktop, wine, and cpanm. Samba was, of course, for the PDC and file/print services. The GUI was needed because a certain software has to be installed on there that needs a GUI. Wine was needed because the software is Windows-native. And cpanm was for a perl script I have running. For Samba, I went into the smb.conf file and enabled domain logons, changed the workgroup/domain name, the logon script for a per-group basis (netlogon/%g), enabled the netlogon and profiles share, and setup a couple of custom shares for the file service. The printer was added later, and seems to be working just fine. I then restarted the services, and used the net groupmap command to ensure my unix groups were mapped correctly to the Windows groups. After this, I went to a Windows box, and was able to successfully join the domain without a problem. After some fidgeting with the software to get it running on the win boxes from the server (it's a records management system program, which stores it's database files on the server), I went to add another computer to the domain. But now it's saying Access Denied. Before when I had this trouble it was because I forgot to add the group "machines" so Samba could create machine accounts. Thinking this was the case, I manually created the machine account to test this theory. However, it would still give me an Access Denied error. That must mean it has something to do with permissions now, correct? I've been fighting with this server for the past two weeks. If it's not one thing that;s wrong, then it's something else completely different. This would be the third time I've actually reinstalled everything to start over. I'll post snippets of my system settings below. If anything else is needed, just say the word and I'll gather up the info. The unix group 'domadmin' is the Domain Admins group. Samba Administrator account administrator:x:1000:1000:Administrator,,,:/home/administrator:/bin/bash Adminstrator's groups administrator adm cdrom sudo dip plugdev lpadmin sambashare domadmin crimestar Samba's Configuration FIle (a snippet anyways) [global] workgroup = CITYPD server string = BPDServer dns proxy = no log file = /var/log/samba/log.%m max log size = 1000 syslog = 0 panic action = /usr/share/samba/panic-action %d security = user encrypt passwords = true passdb backend = tdbsam obey pam restrictions = yes unix password sync = yes passwd program = /usr/bin/passwd %u passwd chat = *Enter\snew\s*\spassword:* %n\n *Retype\snew\s*\spassword:* %n\n *password\supdated\ssuccessfully* . pam password change = yes map to guest = bad user domain logons = yes logon path = \\%L\srv\samba\profiles\%U logon script = logon.bat add machine script = /usr/sbin/useradd -g machines -c "%u machine account" -d /var/lib/samba -s /bin/false %u domain master = yes usershare allow guests = yes [netlogon] comment = Network Logon Service path = /srv/samba/netlogon/%g guest ok = yes read only = yes browseable = no [profiles] comment = All Printers browseable = no path = /var/spool/samba printable = yes guest ok = no read only = yes create mask = 0700 [print$] comment = Printer Drivers path = /var/lib/samba/printers browseable = yes read only = yes guest ok = no write list = root, @lpadmin [crimestar] comment = "Crimestar DB" path = /srv/crimestar/db valid users = @domadmin, @crimestar admin users = administrator writeable = yes guest ok = no browseable = no create mask = 0666 directory mask = 0777 [crimestarfiles] path = /home/administrator/.wine/drive_c/crimestar admin users = administrator browseable = yes ls -la on /srv/samba/profiles drwxrwxrwx 2 root machines 4096 Nov 21 15:27 . drwxr-xr-x 4 root root 4096 Nov 21 15:28 .. ls -la on /srv/samba/netlogon drwxr-xr-x 6 root root 4096 Nov 21 15:30 . drwxr-xr-x 4 root root 4096 Nov 21 15:28 .. drwxr-xr-x 2 root root 4096 Nov 21 15:30 crimestar drwxr-xr-x 2 root root 4096 Nov 21 18:13 domadmin drwxr-xr-x 3 root root 4096 Nov 21 15:30 guests drwxr-xr-x 2 root root 4096 Nov 21 15:29 users GrouMap list Domain Users (S-1-5-21-2978508755-2341913247-928297747-513) -> users Domain Admins (S-1-5-21-2978508755-2341913247-928297747-512) -> domadmin Domain Guests (S-1-5-21-2978508755-2341913247-928297747-514) -> nogroup TLDR I'm getting an Access Denied error message while trying to join a windows box to a samba domain, even after I successfully joined another computer without a problem. System settings / files are quoted above. Anyone have any ideas or suggestions?

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  • nagios NRPE: Unable to read output

    - by user555854
    I currently set up a script to restart my http servers + php5 fpm but can't get it to work. I have googled and have found that mostly permissions are the problems of my error but can't figure it out. I start my script using /usr/lib/nagios/plugins/check_nrpe -H bart -c restart_http This is the output in my syslog on the node I want to restart Jun 27 06:29:35 bart nrpe[8926]: Connection from 192.168.133.17 port 25028 Jun 27 06:29:35 bart nrpe[8926]: Host address is in allowed_hosts Jun 27 06:29:35 bart nrpe[8926]: Handling the connection... Jun 27 06:29:35 bart nrpe[8926]: Host is asking for command 'restart_http' to be run... Jun 27 06:29:35 bart nrpe[8926]: Running command: /usr/bin/sudo /usr/lib/nagios/plugins/http-restart Jun 27 06:29:35 bart nrpe[8926]: Command completed with return code 1 and output: Jun 27 06:29:35 bart nrpe[8926]: Return Code: 1, Output: NRPE: Unable to read output Jun 27 06:29:35 bart nrpe[8926]: Connection from 192.168.133.17 closed. If I run the command myself it runs fine (but asks for a password) (nagios user) This are the script permission and the script contents. -rwxrwxrwx 1 nagios nagios 142 Jun 26 21:41 /usr/lib/nagios/plugins/http-restart #!/bin/bash echo "ok" /etc/init.d/nginx stop /etc/init.d/nginx start /etc/init.d/php5-fpm stop /etc/init.d/php5-fpm start echo "done" I also added this line to visudo nagios ALL=(ALL) NOPASSWD: /usr/lib/nagios/plugins/ My local nagios nrpe.cfg ############################################################################# # Sample NRPE Config File # Written by: Ethan Galstad ([email protected]) # # # NOTES: # This is a sample configuration file for the NRPE daemon. It needs to be # located on the remote host that is running the NRPE daemon, not the host # from which the check_nrpe client is being executed. ############################################################################# # LOG FACILITY # The syslog facility that should be used for logging purposes. log_facility=daemon # PID FILE # The name of the file in which the NRPE daemon should write it's process ID # number. The file is only written if the NRPE daemon is started by the root # user and is running in standalone mode. pid_file=/var/run/nagios/nrpe.pid # PORT NUMBER # Port number we should wait for connections on. # NOTE: This must be a non-priviledged port (i.e. > 1024). # NOTE: This option is ignored if NRPE is running under either inetd or xinetd server_port=5666 # SERVER ADDRESS # Address that nrpe should bind to in case there are more than one interface # and you do not want nrpe to bind on all interfaces. # NOTE: This option is ignored if NRPE is running under either inetd or xinetd #server_address=127.0.0.1 # NRPE USER # This determines the effective user that the NRPE daemon should run as. # You can either supply a username or a UID. # # NOTE: This option is ignored if NRPE is running under either inetd or xinetd nrpe_user=nagios # NRPE GROUP # This determines the effective group that the NRPE daemon should run as. # You can either supply a group name or a GID. # # NOTE: This option is ignored if NRPE is running under either inetd or xinetd nrpe_group=nagios # ALLOWED HOST ADDRESSES # This is an optional comma-delimited list of IP address or hostnames # that are allowed to talk to the NRPE daemon. # # Note: The daemon only does rudimentary checking of the client's IP # address. I would highly recommend adding entries in your /etc/hosts.allow # file to allow only the specified host to connect to the port # you are running this daemon on. # # NOTE: This option is ignored if NRPE is running under either inetd or xinetd allowed_hosts=127.0.0.1,192.168.133.17 # COMMAND ARGUMENT PROCESSING # This option determines whether or not the NRPE daemon will allow clients # to specify arguments to commands that are executed. This option only works # if the daemon was configured with the --enable-command-args configure script # option. # # *** ENABLING THIS OPTION IS A SECURITY RISK! *** # Read the SECURITY file for information on some of the security implications # of enabling this variable. # # Values: 0=do not allow arguments, 1=allow command arguments dont_blame_nrpe=0 # COMMAND PREFIX # This option allows you to prefix all commands with a user-defined string. # A space is automatically added between the specified prefix string and the # command line from the command definition. # # *** THIS EXAMPLE MAY POSE A POTENTIAL SECURITY RISK, SO USE WITH CAUTION! *** # Usage scenario: # Execute restricted commmands using sudo. For this to work, you need to add # the nagios user to your /etc/sudoers. An example entry for alllowing # execution of the plugins from might be: # # nagios ALL=(ALL) NOPASSWD: /usr/lib/nagios/plugins/ # # This lets the nagios user run all commands in that directory (and only them) # without asking for a password. If you do this, make sure you don't give # random users write access to that directory or its contents! command_prefix=/usr/bin/sudo # DEBUGGING OPTION # This option determines whether or not debugging messages are logged to the # syslog facility. # Values: 0=debugging off, 1=debugging on debug=1 # COMMAND TIMEOUT # This specifies the maximum number of seconds that the NRPE daemon will # allow plugins to finish executing before killing them off. command_timeout=60 # CONNECTION TIMEOUT # This specifies the maximum number of seconds that the NRPE daemon will # wait for a connection to be established before exiting. This is sometimes # seen where a network problem stops the SSL being established even though # all network sessions are connected. This causes the nrpe daemons to # accumulate, eating system resources. Do not set this too low. connection_timeout=300 # WEEK RANDOM SEED OPTION # This directive allows you to use SSL even if your system does not have # a /dev/random or /dev/urandom (on purpose or because the necessary patches # were not applied). The random number generator will be seeded from a file # which is either a file pointed to by the environment valiable $RANDFILE # or $HOME/.rnd. If neither exists, the pseudo random number generator will # be initialized and a warning will be issued. # Values: 0=only seed from /dev/[u]random, 1=also seed from weak randomness #allow_weak_random_seed=1 # INCLUDE CONFIG FILE # This directive allows you to include definitions from an external config file. #include=<somefile.cfg> # INCLUDE CONFIG DIRECTORY # This directive allows you to include definitions from config files (with a # .cfg extension) in one or more directories (with recursion). #include_dir=<somedirectory> #include_dir=<someotherdirectory> # COMMAND DEFINITIONS # Command definitions that this daemon will run. Definitions # are in the following format: # # command[<command_name>]=<command_line> # # When the daemon receives a request to return the results of <command_name> # it will execute the command specified by the <command_line> argument. # # Unlike Nagios, the command line cannot contain macros - it must be # typed exactly as it should be executed. # # Note: Any plugins that are used in the command lines must reside # on the machine that this daemon is running on! The examples below # assume that you have plugins installed in a /usr/local/nagios/libexec # directory. Also note that you will have to modify the definitions below # to match the argument format the plugins expect. Remember, these are # examples only! # The following examples use hardcoded command arguments... command[check_users]=/usr/lib/nagios/plugins/check_users -w 5 -c 10 command[check_load]=/usr/lib/nagios/plugins/check_load -w 15,10,5 -c 30,25,20 command[check_hda1]=/usr/lib/nagios/plugins/check_disk -w 20% -c 10% -p /dev/hda1 command[check_zombie_procs]=/usr/lib/nagios/plugins/check_procs -w 5 -c 10 -s Z command[check_total_procs]=/usr/lib/nagios/plugins/check_procs -w 150 -c 200 # The following examples allow user-supplied arguments and can # only be used if the NRPE daemon was compiled with support for # command arguments *AND* the dont_blame_nrpe directive in this # config file is set to '1'. This poses a potential security risk, so # make sure you read the SECURITY file before doing this. #command[check_users]=/usr/lib/nagios/plugins/check_users -w $ARG1$ -c $ARG2$ #command[check_load]=/usr/lib/nagios/plugins/check_load -w $ARG1$ -c $ARG2$ #command[check_disk]=/usr/lib/nagios/plugins/check_disk -w $ARG1$ -c $ARG2$ -p $ARG3$ #command[check_procs]=/usr/lib/nagios/plugins/check_procs -w $ARG1$ -c $ARG2$ -s $ARG3$ command[restart_http]=/usr/lib/nagios/plugins/http-restart # # local configuration: # if you'd prefer, you can instead place directives here include=/etc/nagios/nrpe_local.cfg # # you can place your config snipplets into nrpe.d/ include_dir=/etc/nagios/nrpe.d/ My Sudoers files # /etc/sudoers # # This file MUST be edited with the 'visudo' command as root. # # See the man page for details on how to write a sudoers file. # Defaults env_reset # Host alias specification # User alias specification # Cmnd alias specification # User privilege specification root ALL=(ALL) ALL nagios ALL=(ALL) NOPASSWD: /usr/lib/nagios/plugins/ # Allow members of group sudo to execute any command # (Note that later entries override this, so you might need to move # it further down) %sudo ALL=(ALL) ALL # #includedir /etc/sudoers.d Hopefully someone can help!

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  • Tips on Migrating from AquaLogic .NET Accelerator to WebCenter WSRP Producer for .NET

    - by user647124
    This year I embarked on a journey to migrate a group of ASP.NET web applications developed to integrate with WebLogic Portal 9.2 via the AquaLogic® Interaction .NET Application Accelerator 1.0 to instead use the Oracle WebCenter WSRP Producer for .NET and integrated with WebLogic Portal 10.3.4. It has been a very winding path and this blog entry is intended to share both the lessons learned and relevant approaches that led to those learnings. Like most journeys of discovery, it was not a direct path, and there are notes to let you know when it is practical to skip a section if you are in a hurry to get from here to there. For the Curious From the perspective of necessity, this section would be better at the end. If it were there, though, it would probably be read by far fewer people, including those that are actually interested in these types of sections. Those in a hurry may skip past and be none the worst for it in dealing with the hands-on bits of performing a migration from .NET Accelerator to WSRP Producer. For others who want to talk about why they did what they did after they did it, or just want to know for themselves, enjoy. A Brief (and edited) History of the WSRP for .NET Technologies (as Relevant to the this Post) Note: This section is for those who are curious about why the migration path is not as simple as many other Oracle technologies. You can skip this section in its entirety and still be just as competent in performing a migration as if you had read it. The currently deployed architecture that was to be migrated and upgraded achieved initial integration between .NET and J2EE over the WSRP protocol through the use of The AquaLogic Interaction .NET Application Accelerator. The .NET Accelerator allowed the applications that were written in ASP.NET and deployed on a Microsoft Internet Information Server (IIS) to interact with a WebLogic Portal application deployed on a WebLogic (J2EE application) Server (both version 9.2, the state of the art at the time of its creation). At the time this architectural decision for the application was made, both the AquaLogic and WebLogic brands were owned by BEA Systems. The AquaLogic brand included products acquired by BEA through the acquisition of Plumtree, whose flagship product was a portal platform available in both J2EE and .NET versions. As part of this dual technology support an adaptor was created to facilitate the use of WSRP as a communication protocol where customers wished to integrate components from both versions of the Plumtree portal. The adapter evolved over several product generations to include a broad array of both standard and proprietary WSRP integration capabilities. Later, BEA Systems was acquired by Oracle. Over the course of several years Oracle has acquired a large number of portal applications and has taken the strategic direction to migrate users of these myriad (and formerly competitive) products to the Oracle WebCenter technology stack. As part of Oracle’s strategic technology roadmap, older portal products are being schedule for end of life, including the portal products that were part of the BEA acquisition. The .NET Accelerator has been modified over a very long period of time with features driven by users of that product and developed under three different vendors (each a direct competitor in the same solution space prior to merger). The Oracle WebCenter WSRP Producer for .NET was introduced much more recently with the key objective to specifically address the needs of the WebCenter customers developing solutions accessible through both J2EE and .NET platforms utilizing the WSRP specifications. The Oracle Product Development Team also provides these insights on the drivers for developing the WSRP Producer: ***************************************** Support for ASP.NET AJAX. Controls using the ASP.NET AJAX script manager do not function properly in the Application Accelerator for .NET. Support 2 way SSL in WLP. This was not possible with the proxy/bridge set up in the existing Application Accelerator for .NET. Allow developers to code portlets (Web Parts) using the .NET framework rather than a proprietary framework. Developers had to use the Application Accelerator for .NET plug-ins to Visual Studio to manage preferences and profile data. This is now replaced with the .NET Framework Personalization (for preferences) and Profile providers. The WSRP Producer for .NET was created as a new way of developing .NET portlets. It was never designed to be an upgrade path for the Application Accelerator for .NET. .NET developers would create new .NET portlets with the WSRP Producer for .NET and leave any existing .NET portlets running in the Application Accelerator for .NET. ***************************************** The advantage to creating a new solution for WSRP is a product that is far easier for Oracle to maintain and support which in turn improves quality, reliability and maintainability for their customers. No changes to J2EE applications consuming the WSRP portlets previously rendered by the.NET Accelerator is required to migrate from the Aqualogic WSRP solution. For some customers using the .NET Accelerator the challenge is adapting their current .NET applications to work with the WSRP Producer (or any other WSRP adapter as they are proprietary by nature). Part of this adaptation is the need to deploy the .NET applications as a child to the WSRP producer web application as root. Differences between .NET Accelerator and WSRP Producer Note: This section is for those who are curious about why the migration is not as pluggable as something such as changing security providers in WebLogic Server. You can skip this section in its entirety and still be just as competent in performing a migration as if you had read it. The basic terminology used to describe the participating applications in a WSRP environment are the same when applied to either the .NET Accelerator or the WSRP Producer: Producer and Consumer. In both cases the .NET application serves as what is referred to as a WSRP environment as the Producer. The difference lies in how the two adapters create the WSRP translation of the .NET application. The .NET Accelerator, as the name implies, is meant to serve as a quick way of adding WSRP capability to a .NET application. As such, at a high level, the .NET Accelerator behaves as a proxy for requests between the .NET application and the WSRP Consumer. A WSRP request is sent from the consumer to the .NET Accelerator, the.NET Accelerator transforms this request into an ASP.NET request, receives the response, then transforms the response into a WSRP response. The .NET Accelerator is deployed as a stand-alone application on IIS. The WSRP Producer is deployed as a parent application on IIS and all ASP.NET modules that will be made available over WSRP are deployed as children of the WSRP Producer application. In this manner, the WSRP Producer acts more as a Request Filter than a proxy in the WSRP transactions between Producer and Consumer. Highly Recommended Enabling Logging Note: You can skip this section now, but you will most likely want to come back to it later, so why not just read it now? Logging is very helpful in tracking down the causes of any anomalies during testing of migrated portlets. To enable the WSRP Producer logging, update the Application_Start method in the Global.asax.cs for your .NET application by adding log4net.Config.XmlConfigurator.Configure(); IIS logs will usually (in a standard configuration) be in a sub folder under C:\WINDOWS\system32\LogFiles\W3SVC. WSRP Producer logs will be found at C:\Oracle\Middleware\WSRPProducerForDotNet\wsrpdefault\Logs\WSRPProducer.log InputTrace.webinfo and OutputTrace.webinfo are located under C:\Oracle\Middleware\WSRPProducerForDotNet\wsrpdefault and can be useful in debugging issues related to markup transformations. Things You Must Do Merge Web.Config Note: If you have been skipping all the sections that you can, now is the time to stop and pay attention J Because the existing .NET application will become a sub-application to the WSRP Producer, you will want to merge required settings from the existing Web.Config to the one in the WSRP Producer. Use the WSRP Producer Master Page The Master Page installed for the WSRP Producer provides common, hiddenform fields and JavaScripts to facilitate portlet instance management and display configuration when the child page is being rendered over WSRP. You add the Master Page by including it in the <@ Page declaration with MasterPageFile="~/portlets/Resources/MasterPages/WSRP.Master" . You then replace: <!DOCTYPE HTML PUBLIC "-//W3C//DTD HTML 4.0 Transitional//EN" > <HTML> <HEAD> With <asp:Content ID="ContentHead1" ContentPlaceHolderID="wsrphead" Runat="Server"> And </HEAD> <body> <form id="theForm" method="post" runat="server"> With </asp:Content> <asp:Content ID="ContentBody1" ContentPlaceHolderID="Main" Runat="Server"> And finally </form> </body> </HTML> With </asp:Content> In the event you already use Master Pages, adapt your existing Master Pages to be sub masters. See Nested ASP.NET Master Pages for a detailed reference of how to do this. It Happened to Me, It Might Happen to You…Or Not Watch for Use of Session or Request in OnInit In the event the .NET application being modified has pages developed to assume the user has been authenticated in an earlier page request there may be direct or indirect references in the OnInit method to request or session objects that may not have been created yet. This will vary from application to application, so the recommended approach is to test first. If there is an issue with a page running as a WSRP portlet then check for potential references in the OnInit method (including references by methods called within OnInit) to session or request objects. If there are, the simplest solution is to create a new method and then call that method once the necessary object(s) is fully available. I find doing this at the start of the Page_Load method to be the simplest solution. Case Sensitivity .NET languages are not case sensitive, but Java is. This means it is possible to have many variations of SRC= and src= or .JPG and .jpg. The preferred solution is to make these mark up instances all lower case in your .NET application. This will allow the default Rewriter rules in wsrp-producer.xml to work as is. If this is not practical, then make duplicates of any rules where an issue is occurring due to upper or mixed case usage in the .NET application markup and match the case in use with the duplicate rule. For example: <RewriterRule> <LookFor>(href=\"([^\"]+)</LookFor> <ChangeToAbsolute>true</ChangeToAbsolute> <ApplyTo>.axd,.css</ApplyTo> <MakeResource>true</MakeResource> </RewriterRule> May need to be duplicated as: <RewriterRule> <LookFor>(HREF=\"([^\"]+)</LookFor> <ChangeToAbsolute>true</ChangeToAbsolute> <ApplyTo>.axd,.css</ApplyTo> <MakeResource>true</MakeResource> </RewriterRule> While it is possible to write a regular expression that will handle mixed case usage, it would be long and strenous to test and maintain, so the recommendation is to use duplicate rules. Is it Still Relative? Some .NET applications base relative paths with a fixed root location. With the introduction of the WSRP Producer, the root has moved up one level. References to ~/ will need to be updated to ~/portlets and many ../ paths will need another ../ in front. I Can See You But I Can’t Find You This issue was first discovered while debugging modules with code that referenced the form on a page from the code-behind by name and/or id. The initial error presented itself as run-time error that was difficult to interpret over WSRP but seemed clear when run as straight ASP.NET as it indicated that the object with the form name did not exist. Since the form name was no longer valid after implementing the WSRP Master Page, the likely fix seemed to simply update the references in the code. However, as the WSRP Master Page is external to the code, a compile time error resulted: Error      155         The name 'form1' does not exist in the current context                C:\Oracle\Middleware\WSRPProducerForDotNet\wsrpdefault\portlets\legacywebsite\module\Screens \Reporting.aspx.cs                51           52           legacywebsite.module Much hair-pulling research later it was discovered that it was the use of the FindControl method causing the issue. FindControl doesn’t work quite as expected once a Master Page has been introduced as the controls become embedded in controls, require a recursion to find them that is not part of the FindControl method. In code where the page form is referenced by name, there are two steps to the solution. First, the form needs to be referenced in code generically with Page.Form. For example, this: ToggleControl ctrl = new ToggleControl(frmManualEntry, FunctionLibrary.ParseArrayLst(userObj.Roles)); Becomes this: ToggleControl ctrl = new ToggleControl(Page.Form, FunctionLibrary.ParseArrayLst(userObj.Roles)); Generally the form id is referenced in most ASP.NET applications as a path to a control on the form. To reach the control once a MasterPage has been added requires an additional method to recurse through the controls collections within the form and find the control ID. The following method (found at Rick Strahl's Web Log) corrects this very nicely: public static Control FindControlRecursive(Control Root, string Id) { if (Root.ID == Id) return Root; foreach (Control Ctl in Root.Controls) { Control FoundCtl = FindControlRecursive(Ctl, Id); if (FoundCtl != null) return FoundCtl; } return null; } Where the form name is not referenced, simply using the FindControlRecursive method in place of FindControl will be all that is necessary. Following the second part of the example referenced earlier, the method called with Page.Form changes its value extraction code block from this: Label lblErrMsg = (Label)frmRef.FindControl("lblBRMsg" To this: Label lblErrMsg = (Label) FunctionLibrary.FindControlRecursive(frmRef, "lblBRMsg" The Master That Won’t Step Aside In most migrations it is preferable to make as few changes as possible. In one case I ran across an existing Master Page that would not function as a sub-Master Page. While it would probably have been educational to trace down why, the expedient process of updating it to take the place of the WSRP Master Page is the route I took. The changes are highlighted below: … <asp:ContentPlaceHolder ID="wsrphead" runat="server"></asp:ContentPlaceHolder> </head> <body leftMargin="0" topMargin="0"> <form id="TheForm" runat="server"> <input type="hidden" name="key" id="key" value="" /> <input type="hidden" name="formactionurl" id="formactionurl" value="" /> <input type="hidden" name="handle" id="handle" value="" /> <asp:ScriptManager ID="ScriptManager1" runat="server" EnablePartialRendering="true" > </asp:ScriptManager> This approach did not work for all existing Master Pages, but fortunately all of the other existing Master Pages I have run across worked fine as a sub-Master to the WSRP Master Page. Moving On In Enterprise Portals, even after you get everything working, the work is not finished. Next you need to get it where everyone will work with it. Migration Planning Providing that the server where IIS is running is adequately sized, it is possible to run both the .NET Accelerator and the WSRP Producer on the same server during the upgrade process. The upgrade can be performed incrementally, i.e., one portlet at a time, if server administration processes support it. Those processes would include the ability to manage a second producer in the consuming portal and to change over individual portlet instances from one provider to the other. If processes or requirements demand that all portlets be cut over at the same time, it needs to be determined if this cut over should include a new producer, updating all of the portlets in the consumer, or if the WSRP Producer portlet configuration must maintain the naming conventions used by the .NET Accelerator and simply change the WSRP end point configured in the consumer. In some enterprises it may even be necessary to maintain the same WSDL end point, at which point the IIS configuration will be where the updates occur. The downside to such a requirement is that it makes rolling back very difficult, should the need arise. Location, Location, Location Not everyone wants the web application to have the descriptively obvious wsrpdefault location, or needs to create a second WSRP site on the same server. The instructions below are from the product team and, while targeted towards making a second site, will work for creating a site with a different name and then remove the old site. You can also change just the name in IIS. Manually Creating a WSRP Producer Site Instructions (NOTE: all executables used are the same ones used by the installer and “wsrpdev” will be the name of the new instance): 1. Copy C:\Oracle\Middleware\WSRPProducerForDotNet\wsrpdefault to C:\Oracle\Middleware\WSRPProducerForDotNet\wsrpdev. 2. Bring up a command window as an administrator 3. Run C:\Oracle\Middleware\WSRPProducerForDotNet\uninstall_resources\IISAppAccelSiteCreator.exe install WSRPProducers wsrpdev "C:\Oracle\Middleware\WSRPProducerForDotNet\wsrpdev" 8678 2.0.50727 4. Run C:\Oracle\Middleware\WSRPProducerForDotNet\uninstall_resources\PermManage.exe add FileSystem C:\Oracle\Middleware\WSRPProducerForDotNet\wsrpdev "NETWORK SERVICE" 3 1 5. Run C:\Oracle\Middleware\WSRPProducerForDotNet\uninstall_resources\PermManage.exe add FileSystem C:\Oracle\Middleware\WSRPProducerForDotNet\wsrpdev EVERYONE 1 1 6. Open up C:\Oracle\Middleware\WSRPProducerForDotNet\wsdl\1.0\WSRPService.wsdl and replace wsrpdefault with wsrpdev 7. Open up C:\Oracle\Middleware\WSRPProducerForDotNet\wsdl\2.0\WSRPService.wsdl and replace wsrpdefault with wsrpdev Tests: 1. Bring up a browser on the host itself and go to http://localhost:8678/wsrpdev/wsdl/1.0/WSRPService.wsdl and make sure that the URLs in the XML returned include the wsrpdev changes you made in step 6. 2. Bring up a browser on the host itself and see if the default sample comes up: http://localhost:8678/wsrpdev/portlets/ASPNET_AJAX_sample/default.aspx 3. Register the producer in WLP and test the portlet. Changing the Port used by WSRP Producer The pre-configured port for the WSRP Producer is 8678. You can change this port by updating both the IIS configuration and C:\Oracle\Middleware\WSRPProducerForDotNet\[WSRP_APP_NAME]\wsdl\1.0\WSRPService.wsdl. Do You Need to Migrate? Oracle Premier Support ended in November of 2010 for AquaLogic Interaction .NET Application Accelerator 1.x and Extended Support ends in November 2012 (see http://www.oracle.com/us/support/lifetime-support/lifetime-support-software-342730.html for other related dates). This means that integration with products released after November of 2010 is not supported. If having such support is the policy within your enterprise, you do indeed need to migrate. If changes in your enterprise cause your current solution with the .NET Accelerator to no longer function properly, you may need to migrate. Migration is a choice, and if the goals of your enterprise are to take full advantage of newer technologies then migration is certainly one activity you should be planning for.

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  • Informed TDD &ndash; Kata &ldquo;To Roman Numerals&rdquo;

    - by Ralf Westphal
    Originally posted on: http://geekswithblogs.net/theArchitectsNapkin/archive/2014/05/28/informed-tdd-ndash-kata-ldquoto-roman-numeralsrdquo.aspxIn a comment on my article on what I call Informed TDD (ITDD) reader gustav asked how this approach would apply to the kata “To Roman Numerals”. And whether ITDD wasn´t a violation of TDD´s principle of leaving out “advanced topics like mocks”. I like to respond with this article to his questions. There´s more to say than fits into a commentary. Mocks and TDD I don´t see in how far TDD is avoiding or opposed to mocks. TDD and mocks are orthogonal. TDD is about pocess, mocks are about structure and costs. Maybe by moving forward in tiny red+green+refactor steps less need arises for mocks. But then… if the functionality you need to implement requires “expensive” resource access you can´t avoid using mocks. Because you don´t want to constantly run all your tests against the real resource. True, in ITDD mocks seem to be in almost inflationary use. That´s not what you usually see in TDD demonstrations. However, there´s a reason for that as I tried to explain. I don´t use mocks as proxies for “expensive” resource. Rather they are stand-ins for functionality not yet implemented. They allow me to get a test green on a high level of abstraction. That way I can move forward in a top-down fashion. But if you think of mocks as “advanced” or if you don´t want to use a tool like JustMock, then you don´t need to use mocks. You just need to stand the sight of red tests for a little longer ;-) Let me show you what I mean by that by doing a kata. ITDD for “To Roman Numerals” gustav asked for the kata “To Roman Numerals”. I won´t explain the requirements again. You can find descriptions and TDD demonstrations all over the internet, like this one from Corey Haines. Now here is, how I would do this kata differently. 1. Analyse A demonstration of TDD should never skip the analysis phase. It should be made explicit. The requirements should be formalized and acceptance test cases should be compiled. “Formalization” in this case to me means describing the API of the required functionality. “[D]esign a program to work with Roman numerals” like written in this “requirement document” is not enough to start software development. Coding should only begin, if the interface between the “system under development” and its context is clear. If this interface is not readily recognizable from the requirements, it has to be developed first. Exploration of interface alternatives might be in order. It might be necessary to show several interface mock-ups to the customer – even if that´s you fellow developer. Designing the interface is a task of it´s own. It should not be mixed with implementing the required functionality behind the interface. Unfortunately, though, this happens quite often in TDD demonstrations. TDD is used to explore the API and implement it at the same time. To me that´s a violation of the Single Responsibility Principle (SRP) which not only should hold for software functional units but also for tasks or activities. In the case of this kata the API fortunately is obvious. Just one function is needed: string ToRoman(int arabic). And it lives in a class ArabicRomanConversions. Now what about acceptance test cases? There are hardly any stated in the kata descriptions. Roman numerals are explained, but no specific test cases from the point of view of a customer. So I just “invent” some acceptance test cases by picking roman numerals from a wikipedia article. They are supposed to be just “typical examples” without special meaning. Given the acceptance test cases I then try to develop an understanding of the problem domain. I´ll spare you that. The domain is trivial and is explain in almost all kata descriptions. How roman numerals are built is not difficult to understand. What´s more difficult, though, might be to find an efficient solution to convert into them automatically. 2. Solve The usual TDD demonstration skips a solution finding phase. Like the interface exploration it´s mixed in with the implementation. But I don´t think this is how it should be done. I even think this is not how it really works for the people demonstrating TDD. They´re simplifying their true software development process because they want to show a streamlined TDD process. I doubt this is helping anybody. Before you code you better have a plan what to code. This does not mean you have to do “Big Design Up-Front”. It just means: Have a clear picture of the logical solution in your head before you start to build a physical solution (code). Evidently such a solution can only be as good as your understanding of the problem. If that´s limited your solution will be limited, too. Fortunately, in the case of this kata your understanding does not need to be limited. Thus the logical solution does not need to be limited or preliminary or tentative. That does not mean you need to know every line of code in advance. It just means you know the rough structure of your implementation beforehand. Because it should mirror the process described by the logical or conceptual solution. Here´s my solution approach: The arabic “encoding” of numbers represents them as an ordered set of powers of 10. Each digit is a factor to multiply a power of ten with. The “encoding” 123 is the short form for a set like this: {1*10^2, 2*10^1, 3*10^0}. And the number is the sum of the set members. The roman “encoding” is different. There is no base (like 10 for arabic numbers), there are just digits of different value, and they have to be written in descending order. The “encoding” XVI is short for [10, 5, 1]. And the number is still the sum of the members of this list. The roman “encoding” thus is simpler than the arabic. Each “digit” can be taken at face value. No multiplication with a base required. But what about IV which looks like a contradiction to the above rule? It is not – if you accept roman “digits” not to be limited to be single characters only. Usually I, V, X, L, C, D, M are viewed as “digits”, and IV, IX etc. are viewed as nuisances preventing a simple solution. All looks different, though, once IV, IX etc. are taken as “digits”. Then MCMLIV is just a sum: M+CM+L+IV which is 1000+900+50+4. Whereas before it would have been understood as M-C+M+L-I+V – which is more difficult because here some “digits” get subtracted. Here´s the list of roman “digits” with their values: {1, I}, {4, IV}, {5, V}, {9, IX}, {10, X}, {40, XL}, {50, L}, {90, XC}, {100, C}, {400, CD}, {500, D}, {900, CM}, {1000, M} Since I take IV, IX etc. as “digits” translating an arabic number becomes trivial. I just need to find the values of the roman “digits” making up the number, e.g. 1954 is made up of 1000, 900, 50, and 4. I call those “digits” factors. If I move from the highest factor (M=1000) to the lowest (I=1) then translation is a two phase process: Find all the factors Translate the factors found Compile the roman representation Translation is just a look-up. Finding, though, needs some calculation: Find the highest remaining factor fitting in the value Remember and subtract it from the value Repeat with remaining value and remaining factors Please note: This is just an algorithm. It´s not code, even though it might be close. Being so close to code in my solution approach is due to the triviality of the problem. In more realistic examples the conceptual solution would be on a higher level of abstraction. With this solution in hand I finally can do what TDD advocates: find and prioritize test cases. As I can see from the small process description above, there are two aspects to test: Test the translation Test the compilation Test finding the factors Testing the translation primarily means to check if the map of factors and digits is comprehensive. That´s simple, even though it might be tedious. Testing the compilation is trivial. Testing factor finding, though, is a tad more complicated. I can think of several steps: First check, if an arabic number equal to a factor is processed correctly (e.g. 1000=M). Then check if an arabic number consisting of two consecutive factors (e.g. 1900=[M,CM]) is processed correctly. Then check, if a number consisting of the same factor twice is processed correctly (e.g. 2000=[M,M]). Finally check, if an arabic number consisting of non-consecutive factors (e.g. 1400=[M,CD]) is processed correctly. I feel I can start an implementation now. If something becomes more complicated than expected I can slow down and repeat this process. 3. Implement First I write a test for the acceptance test cases. It´s red because there´s no implementation even of the API. That´s in conformance with “TDD lore”, I´d say: Next I implement the API: The acceptance test now is formally correct, but still red of course. This will not change even now that I zoom in. Because my goal is not to most quickly satisfy these tests, but to implement my solution in a stepwise manner. That I do by “faking” it: I just “assume” three functions to represent the transformation process of my solution: My hypothesis is that those three functions in conjunction produce correct results on the API-level. I just have to implement them correctly. That´s what I´m trying now – one by one. I start with a simple “detail function”: Translate(). And I start with all the test cases in the obvious equivalence partition: As you can see I dare to test a private method. Yes. That´s a white box test. But as you´ll see it won´t make my tests brittle. It serves a purpose right here and now: it lets me focus on getting one aspect of my solution right. Here´s the implementation to satisfy the test: It´s as simple as possible. Right how TDD wants me to do it: KISS. Now for the second equivalence partition: translating multiple factors. (It´a pattern: if you need to do something repeatedly separate the tests for doing it once and doing it multiple times.) In this partition I just need a single test case, I guess. Stepping up from a single translation to multiple translations is no rocket science: Usually I would have implemented the final code right away. Splitting it in two steps is just for “educational purposes” here. How small your implementation steps are is a matter of your programming competency. Some “see” the final code right away before their mental eye – others need to work their way towards it. Having two tests I find more important. Now for the next low hanging fruit: compilation. It´s even simpler than translation. A single test is enough, I guess. And normally I would not even have bothered to write that one, because the implementation is so simple. I don´t need to test .NET framework functionality. But again: if it serves the educational purpose… Finally the most complicated part of the solution: finding the factors. There are several equivalence partitions. But still I decide to write just a single test, since the structure of the test data is the same for all partitions: Again, I´m faking the implementation first: I focus on just the first test case. No looping yet. Faking lets me stay on a high level of abstraction. I can write down the implementation of the solution without bothering myself with details of how to actually accomplish the feat. That´s left for a drill down with a test of the fake function: There are two main equivalence partitions, I guess: either the first factor is appropriate or some next. The implementation seems easy. Both test cases are green. (Of course this only works on the premise that there´s always a matching factor. Which is the case since the smallest factor is 1.) And the first of the equivalence partitions on the higher level also is satisfied: Great, I can move on. Now for more than a single factor: Interestingly not just one test becomes green now, but all of them. Great! You might say, then I must have done not the simplest thing possible. And I would reply: I don´t care. I did the most obvious thing. But I also find this loop very simple. Even simpler than a recursion of which I had thought briefly during the problem solving phase. And by the way: Also the acceptance tests went green: Mission accomplished. At least functionality wise. Now I´ve to tidy up things a bit. TDD calls for refactoring. Not uch refactoring is needed, because I wrote the code in top-down fashion. I faked it until I made it. I endured red tests on higher levels while lower levels weren´t perfected yet. But this way I saved myself from refactoring tediousness. At the end, though, some refactoring is required. But maybe in a different way than you would expect. That´s why I rather call it “cleanup”. First I remove duplication. There are two places where factors are defined: in Translate() and in Find_factors(). So I factor the map out into a class constant. Which leads to a small conversion in Find_factors(): And now for the big cleanup: I remove all tests of private methods. They are scaffolding tests to me. They only have temporary value. They are brittle. Only acceptance tests need to remain. However, I carry over the single “digit” tests from Translate() to the acceptance test. I find them valuable to keep, since the other acceptance tests only exercise a subset of all roman “digits”. This then is my final test class: And this is the final production code: Test coverage as reported by NCrunch is 100%: Reflexion Is this the smallest possible code base for this kata? Sure not. You´ll find more concise solutions on the internet. But LOC are of relatively little concern – as long as I can understand the code quickly. So called “elegant” code, however, often is not easy to understand. The same goes for KISS code – especially if left unrefactored, as it is often the case. That´s why I progressed from requirements to final code the way I did. I first understood and solved the problem on a conceptual level. Then I implemented it top down according to my design. I also could have implemented it bottom-up, since I knew some bottom of the solution. That´s the leaves of the functional decomposition tree. Where things became fuzzy, since the design did not cover any more details as with Find_factors(), I repeated the process in the small, so to speak: fake some top level, endure red high level tests, while first solving a simpler problem. Using scaffolding tests (to be thrown away at the end) brought two advantages: Encapsulation of the implementation details was not compromised. Naturally private methods could stay private. I did not need to make them internal or public just to be able to test them. I was able to write focused tests for small aspects of the solution. No need to test everything through the solution root, the API. The bottom line thus for me is: Informed TDD produces cleaner code in a systematic way. It conforms to core principles of programming: Single Responsibility Principle and/or Separation of Concerns. Distinct roles in development – being a researcher, being an engineer, being a craftsman – are represented as different phases. First find what, what there is. Then devise a solution. Then code the solution, manifest the solution in code. Writing tests first is a good practice. But it should not be taken dogmatic. And above all it should not be overloaded with purposes. And finally: moving from top to bottom through a design produces refactored code right away. Clean code thus almost is inevitable – and not left to a refactoring step at the end which is skipped often for different reasons.   PS: Yes, I have done this kata several times. But that has only an impact on the time needed for phases 1 and 2. I won´t skip them because of that. And there are no shortcuts during implementation because of that.

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  • how to install ffmpeg in cpanel

    - by Ajay Chthri
    i'm using dedicated server(linux) so i need to install ffmpeg in cpanel so here ffmpeg i found in Main Software Install a Perl Module but i writing script in php so how can i install ffmpeg phpperl when i'am trying to install ffmpeg in perl module i get this response Checking C compiler....C compiler (/usr/bin/cc) OK (cached Tue Jan 17 19:16:31 2012)....Done CPAN fallback is disabled since /var/cpanel/conserve_memory exists, and cpanm is available. Method: Using Perl Expect, Installer: cpanm You have make /usr/bin/make Falling back to HTTP::Tiny 0.009 You have /bin/tar: tar (GNU tar) 1.15.1 You have /usr/bin/unzip You have Cpanel::HttpRequest 2.1 Testing connection speed...(using fast method)...Done Ping:2 (ticks) Testing connection speed to cpan.knowledgematters.net using pureperl...(28800.00 bytes/s)...Done Ping:2 (ticks) Testing connection speed to cpan.develooper.com using pureperl...(22233.33 bytes/s)...Done Ping:2 (ticks) Testing connection speed to cpan.schatt.com using pureperl...(32750.00 bytes/s)...Done Ping:3 (ticks) Testing connection speed to cpan.mirror.facebook.net using pureperl...(14050.00 bytes/s)...Done Ping:2 (ticks) Testing connection speed to cpan.mirrors.hoobly.com using pureperl...(5150.00 bytes/s)...Done Five usable mirrors located Ping:0 (ticks) Testing connection speed to 208.109.109.239 using pureperl...(28950.00 bytes/s)...Done Ping:2 (ticks) Testing connection speed to 208.82.118.100 using pureperl...(19300.00 bytes/s)...Done Ping:1 (ticks) Testing connection speed to 69.50.192.73 using pureperl...(19300.00 bytes/s)...Done Three usable fallback mirrors located Mirror Check passed for cpan.schatt.com (/index.html) Searching on cpanmetadb ... Fetching http://cpanmetadb.cpanel.net/v1.0/package/Video::FFmpeg?cpanel_version=11.30.5.6&cpanel_tier=release (connected:0).......(request attempt 1/12)...Using dns cache file /root/.HttpRequest/cpanmetadb.cpanel.net......searching for mirrors (mirror search attempt 1/3)......5 usable mirrors located. (less then expected)......mirror search success......connecting to 208.74.123.82...@208.74.123.82......connected......receiving...100%......request success......Done Searching Video::FFmpeg on cpanmetadb (http://cpanmetadb.cpanel.net/v1.0/package/Video::FFmpeg?cpanel_version=11.30.5.6&cpanel_tier=release) ... Fetching http://cpanmetadb.cpanel.net/v1.0/package/Video::FFmpeg?cpanel_version=11.30.5.6&cpanel_tier=release (connected:1).......(request attempt 1/12)[email protected]%......request success......Done Source: fastest CPAN mirror ... --> Working on Video::FFmpeg Fetching http://cpan.schatt.com//authors/id/R/RA/RANDOMMAN/Video-FFmpeg-0.47.tar.gz ... Fetching http://cpan.schatt.com/authors/id/R/RA/RANDOMMAN/Video-FFmpeg-0.47.tar.gz (connected:1).......(request attempt 1/12)...Resolving cpan.schatt.com...(resolve attempt 1/65)......connecting to 66.249.128.125...@66.249.128.125......connected......receiving...25%...50%...75%...100%......request success......Done OK Unpacking Video-FFmpeg-0.47.tar.gz Video-FFmpeg-0.47/ Video-FFmpeg-0.47/Changes Video-FFmpeg-0.47/FFmpeg.xs Video-FFmpeg-0.47/MANIFEST Video-FFmpeg-0.47/META.yml Video-FFmpeg-0.47/Makefile.PL Video-FFmpeg-0.47/README Video-FFmpeg-0.47/lib/ Video-FFmpeg-0.47/lib/Video/ Video-FFmpeg-0.47/lib/Video/FFmpeg/ Video-FFmpeg-0.47/lib/Video/FFmpeg/AVFormat.pm Video-FFmpeg-0.47/lib/Video/FFmpeg/AVStream/ Video-FFmpeg-0.47/lib/Video/FFmpeg/AVStream/Audio.pm Video-FFmpeg-0.47/lib/Video/FFmpeg/AVStream/Subtitle.pm Video-FFmpeg-0.47/lib/Video/FFmpeg/AVStream/Video.pm Video-FFmpeg-0.47/lib/Video/FFmpeg/AVStream.pm Video-FFmpeg-0.47/lib/Video/FFmpeg.pm Video-FFmpeg-0.47/ppport.h Video-FFmpeg-0.47/t/ Video-FFmpeg-0.47/t/Video-FFmpeg.t Video-FFmpeg-0.47/test Video-FFmpeg-0.47/test.mp4 Video-FFmpeg-0.47/typemap Entering Video-FFmpeg-0.47 Checking configure dependencies from META.yml META.yml not found or unparsable. Fetching META.yml from search.cpan.org Fetching http://search.cpan.org/meta/Video-FFmpeg-0.47/META.yml (connected:1).......(request attempt 1/12)...Resolving search.cpan.org...(resolve attempt 1/65)......connecting to 199.15.176.161...@199.15.176.161......connected......receiving...100%......request success......Done Configuring Video-FFmpeg-0.47 ... Running Makefile.PL Perl v5.10.0 required--this is only v5.8.8, stopped at Makefile.PL line 1. BEGIN failed--compilation aborted at Makefile.PL line 1. N/A ! Configure failed for Video-FFmpeg-0.47. See /home/.cpanm/build.log for details. Perl Expect failed with non-zero exit status: 256 All available perl module install methods have failed guide me how can i install ffmpeg in cPanel Thanks for advance.

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  • Weird networking problem ( Linksys, Windows 7 )

    - by Rohit Nair
    Okay it's a bit tough to figure out where to start from, but here is the basic summary of the issue: During general internet usage, there are times when any attempt to visit a website stalls at "Waiting for somedomain.com". This problem occurs in Firefox, IE and Chrome. No website will load, INCLUDING the router configuration page at 192.168.1.1. Curiously, ping works fine, and other network apps such as MSN Messenger continue to work and I can send and receive messages. Disconnecting and reconnecting to the wireless network seems to fix the problem for a bit, but there are times when it relapses into not loading after every 2-3 http requests. Restarting the router seems to fix the issue, but it can crop up hours or days later. I have a CCNA cert and I know my way around the Windows family of operating systems, so I'm going to list all the things I've tried here. Other computers on the network seem to suffer the same problem, which makes me think it might be a specific problem with something in Win7. The random nature of this issue makes it a bit difficult to confirm, but I can definitely say that I have experienced this on the following systems: Windows 7 64-bit on my desktop Windows Vista 32-bit on my desktop ( the desktop has 2 wireless NICs and the problem existed on both ) Windows Vista 32-bit on my laptop ( both with wireless and wired ) Windows XP SP3 on another laptop ( both wireless and wired ) Using Wireshark to sniff packets seemed to indicate that although HTTP requests were being SENT out, no packets were coming in to respond to the HTTP request. However, other network apps continued to work i.e I would still receive IMs on Windows Live Messenger. Disabling IPV6 had no effect. Updating router firmware to the latest stock firmware by Linksys had no effect. Switching to dd-wrt firmware had no effect. By "no effect" I mean that although the restart required by firmware updates fixed the problem at the time, it still came back. A couple of weeks back, after a LOT of googling and flipping of various options, I figured it might be a case of router slowdown ( http://www.dd-wrt.com/wiki/index.php/Router%5FSlowdown ) caused by the fact that I occasionally run a torrent client. I tried changing the configuration as suggested in that router slowdown link, and restarted the router. However I have not run the torrent client for 12 days now, and yet I still randomly experience this problem. Currently the computer I am using is running Windows 7 64-bit. I would just like to reiterate some of the reasons that I was confused by the issue. Even the router config page at 192.168.1.1 would not load, indicating that it's not a problem with the WAN link, but probably a router issue or a local computer issue. For some reason, disconnecting and reconnecting to the wireless network immediately seems to fix the problem. Updating the router firmware, even switching to open source firmware did nothing. So it seemed to be a computer issue. On the other hand, I have not seen any mass outrage of people having networking problems with Windows 7 and Linksys routers, especially a problem of this sort, and I have tweaked every network setting I could think of. Although HTTP seems to have trouble, ping works fine, DNS lookups work fine, other networking apps work fine. However if I disconnect from Windows Live Messenger and try to reconnect, it fails to reconnect. So although it could receive data over the existing TCP/IP connection, trying to start a new one failed? Does anyone have any further ideas on debugging or fixing this issue? I am reasonably certain there are no viruses or other malicious apps on my network, and I am also reasonably certain that nobody is accessing my router without my consent. Router: Linksys WRT54G2 1.0 running dd-wrt firmware Wireless Card: Alfa AWUS036H OS: Windows 7 64-bit EDIT: I tried switching to a clean wireless channel free from interference, but the problem still persisted. I tried connecting directly with a cable, but the problem still persisted. Signed A very confused and bewildered geek whose knowledge seems to be useless in the face of this frustrating network issue.

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  • Persistent static routes fail on MacOS 10.6.5 startup!

    - by verbalicious
    I'm unable to get static routes to persist a reboot on Mac OS 10.6.5. I've tried all of the methods prescribed in Google search results, and previous posts on this site. I've tried manually creating a launchd daemon, and used RouteSplit's launchd daemon to no avail. It's clear that the interface is not ready when these methods attempt to apply the route. This workstation in question is getting its IP from DHCP and probably hasn't gotten its DHCP lease when the command runs. We're able to apply the route by hand when logged in, but not through startup methods. Is there another way to apply this route by sneaking the command into something later, but before the login window appears to the user? Here is some relevant log info from system.log. You can see the "route: writing to routing socket: Network is unreachable" errors where my launchd script fires off. I've tried adding extra "sleep" and "ipconfig waitall" statements later in the script but this doesn't fly. Dec 15 19:30:41 localhost com.apple.launchd[1]: *** launchd[1] has started up. *** Dec 15 19:30:45 localhost mDNSResponder[18]: mDNSResponder mDNSResponder-258.13 (Oct 8 2010 17:10:30) starting Dec 15 19:30:47 localhost configd[15]: bootp_session_transmit: bpf_write(en1) failed: Network is down (50) Dec 15 19:30:47 localhost configd[15]: DHCP en1: INIT transmit failed Dec 15 19:30:47 localhost configd[15]: network configuration changed. Dec 15 19:30:47 Administrators-MacBook-Pro configd[15]: setting hostname to "Administrators-MacBook-Pro.local" Dec 15 19:30:47 Administrators-MacBook-Pro blued[16]: Apple Bluetooth daemon started Dec 15 19:30:52 Administrators-MacBook-Pro syslog[67]: routes.sh: Starting RouteSplit Dec 15 19:30:53 Administrators-MacBook-Pro com.apple.usbmuxd[41]: usbmuxd-207 built for iTunesTenOne on Oct 19 2010 at 13:50:35, running 64 bit Dec 15 19:30:54 Administrators-MacBook-Pro /System/Library/CoreServices/loginwindow.app/Contents/MacOS/loginwindow[50]: Login Window Application Started Dec 15 19:30:55 Administrators-MacBook-Pro bootlog[61]: BOOT_TIME: 1292459441 0 Dec 15 19:30:55 Administrators-MacBook-Pro syslog[86]: routes.sh: static route 192.168.0.0/23 192.168.2.2 Dec 15 19:30:55 Administrators-MacBook-Pro net.routes.static[65]: route: writing to routing socket: Network is unreachable Dec 15 19:30:55 Administrators-MacBook-Pro net.routes.static[65]: add net 192.168.0.0: gateway 192.168.2.2: Network is unreachable Dec 15 19:30:57 Administrators-MacBook-Pro org.apache.httpd[38]: httpd: Could not reliably determine the server's fully qualified domain name, using Administrators-MacBook-Pro.local for ServerName Dec 15 19:30:58 Administrators-MacBook-Pro loginwindow[50]: Login Window Started Security Agent Dec 15 19:30:58 Administrators-MacBook-Pro WindowServer[89]: kCGErrorFailure: Set a breakpoint @ CGErrorBreakpoint() to catch errors as they are logged. Dec 15 19:30:58 Administrators-MacBook-Pro com.apple.WindowServer[89]: Wed Dec 15 19:30:58 Administrators-MacBook-Pro.local WindowServer[89] <Error>: kCGErrorFailure: Set a breakpoint @ CGErrorBreakpoint() to catch errors as they are logged. Dec 15 19:31:18 Administrators-MacBook-Pro configd[15]: network configuration changed. Dec 15 19:31:19 administrators-macbook-pro configd[15]: setting hostname to "administrators-macbook-pro.local" Dec 15 19:31:25 administrators-macbook-pro _mdnsresponder[121]: /usr/libexec/ntpd-wrapper: scutil key State:/Network/Global/DNS not present after 30 seconds Dec 15 19:31:25 administrators-macbook-pro _mdnsresponder[124]: sntp options: a=2 v=1 e=0.100 E=5.000 P=2147483647.000 Dec 15 19:31:25 administrators-macbook-pro _mdnsresponder[124]: d=15 c=5 x=0 op=1 l=/var/run/sntp.pid f= time.apple.com Dec 15 19:31:25 administrators-macbook-pro _mdnsresponder[124]: sntp: getaddrinfo(hostname, ntp) failed with nodename nor servname provided, or not known Dec 15 19:31:27 administrators-macbook-pro configd[15]: network configuration changed. Dec 15 19:31:27 Administrators-MacBook-Pro configd[15]: setting hostname to "Administrators-MacBook-Pro.local" Dec 15 19:31:27 Administrators-MacBook-Pro ntpd[37]: Cannot find existing interface for address 17.151.16.20 Dec 15 19:31:27 Administrators-MacBook-Pro ntpd_initres[125]: ntpd indicates no data available! Dec 15 19:31:31 Administrators-MacBook-Pro sshd[128]: USER_PROCESS: 133 ttys000 Dec 15 19:31:37 Administrators-MacBook-Pro sudo[138]: administrator : TTY=ttys000 ; PWD=/Users/administrator ; USER=root ; COMMAND=/usr/bin/less /var/log/system.log ``You can see the following line in /var/log/kernel.log that shows the en0 interface coming up: Dec 15 19:30:51 Administrators-MacBook-Pro kernel[0]: Ethernet [AppleBCM5701Ethernet]: Link up on en0, 1-Gigabit, Full-duplex, No flow-control, Debug [796d,0f01,0de1,0300,c1e1,3800]

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  • GMail suspects confirmation email in stealing personal information

    - by Dennis Gorelik
    When user registers on my web site, web site sends user email confirmation link. Subject: Please confirm your email address Body:Please open this link in your browser to confirm your email address: http://www.postjobfree.com/a/c301718062444f96ba0e358ea833c9b3 This link will expire on: 6/9/2012 8:04:07 PM EST. If my web site sends that email to GMaill (either @gmail.com or another domain that's handled by Google Apps) and that user never emailed to email -- then GMail not only puts the email to spam folder, but also adds prominent red warning:Be careful with this message. Similar messages were used to steal people's personal information. Unless you trust the sender, don't click links or reply with personal information. Learn more That warning really scares many of my users, so they are afraid to open that link and confirm their email. What can I do about it? Ideally I would like that message end up in user's inbox, not spam folder. But at least how do I prevent that scary message? IP address of my mailing server is not blacklisted: http://www.mxtoolbox.com/SuperTool.aspx?action=blacklist%3a208.43.198.72 I use SPF and DKIM signature. Below is the email that ended up in spam folder with that scary red message. Delivered-To: [email protected] Received: by 10.112.84.98 with SMTP id x2csp36568lby; Fri, 8 Jun 2012 17:04:15 -0700 (PDT) Received: by 10.60.25.6 with SMTP id y6mr9110318oef.42.1339200255375; Fri, 08 Jun 2012 17:04:15 -0700 (PDT) Return-Path: Received: from smtp.postjobfree.com (smtp.postjobfree.com. [208.43.198.72]) by mx.google.com with ESMTP id v8si6058193oev.44.2012.06.08.17.04.14; Fri, 08 Jun 2012 17:04:15 -0700 (PDT) Received-SPF: pass (google.com: domain of [email protected] designates 208.43.198.72 as permitted sender) client-ip=208.43.198.72; Authentication-Results: mx.google.com; spf=pass (google.com: domain of [email protected] designates 208.43.198.72 as permitted sender) [email protected]; dkim=pass [email protected] DomainKey-Signature: a=rsa-sha1; c=nofws; q=dns; d=postjobfree.com; s=postjobfree.com; h= received:message-id:mime-version:from:to:date:subject:content-type; b=TCip/3hP1WWViWB1cdAzMFPjyi/aUKXQbuSTVpEO7qr8x3WdMFhJCqZciA69S0HB4 Koatk2cQQ3fOilr4ledCgZYemLSJgwa/ZRhObnqgPHAglkBy8/RAwkrwaE0GjLKup 0XI6G2wPlh+ReR+inkMwhCPHFInmvrh4evlBx/VlA= DKIM-Signature: v=1; a=rsa-sha256; c=relaxed/relaxed; d=postjobfree.com; s=postjobfree.com; h=content-type:subject:date:to:from:mime-version:message-id; bh=N59EIgRECIlAnd41LY4HY/OFI+v1p7t5M9yP+3FsKXY=; b=J3/BdZmpjzP4I6GA4ntmi4REu5PpOcmyzEL+6i7y7LaTR8tuc2h7fdW4HaMPlB7za Lj4NJPed61ErumO66eG4urd1UfyaRDtszWeuIbcIUqzwYpnMZ8ytaj8DPcWPE3JYj oKhcYyiVbgiFjLujib3/2k2PqDIrNutRH9Ln7puz4= Received: from sv3035 (sv3035 [208.43.198.72]) by smtp.postjobfree.com with SMTP; Fri, 8 Jun 2012 20:04:07 -0400 Message-ID: MIME-Version: 1.0 From: "PostJobFree Notification" To: [email protected] Date: 8 Jun 2012 20:04:07 -0400 Subject: Please confirm your email address Content-Type: multipart/alternative; boundary=--boundary_107_ffa6a9ea-01dc-40f5-a50c-4c3b3d113f08 ----boundary_107_ffa6a9ea-01dc-40f5-a50c-4c3b3d113f08 Content-Type: text/plain; charset=us-ascii Content-Transfer-Encoding: quoted-printable Please open this link in your browser to confirm your email addre= ss: =0D=0Ahttp://www.postjobfree.com/a/c301718062444f96ba0e358ea8= 33c9b3 =0D=0AThis link will expire on: 6/9/2012 8:04:07 PM EST. =0D=0A ----boundary_107_ffa6a9ea-01dc-40f5-a50c-4c3b3d113f08 Content-Type: text/html; charset=utf-8 Content-Transfer-Encoding: base64 PGh0bWw+PGhlYWQ+PG1ldGEgaHR0cC1lcXVpdj1Db250ZW50LVR5cGUgY29udGVu dD0idGV4dC9odG1sOyBjaGFyc2V0PXV0Zi04Ij48L2hlYWQ+DQo8Ym9keT48ZGl2 Pg0KUGxlYXNlIG9wZW4gdGhpcyBsaW5rIGluIHlvdXIgYnJvd3NlciB0byBjb25m aXJtIHlvdXIgZW1haWwgYWRkcmVzczo8YnIgLz48YSBocmVmPSJodHRwOi8vd3d3 LnBvc3Rqb2JmcmVlLmNvbS9hL2MzMDE3MTgwNjI0NDRmOTZiYTBlMzU4ZWE4MzNj OWIzIj5odHRwOi8vd3d3LnBvc3Rqb2JmcmVlLmNvbS9hL2MzMDE3MTgwNjI0NDRm OTZiYTBlMzU4ZWE4MzNjOWIzPC9hPjxiciAvPlRoaXMgbGluayB3aWxsIGV4cGly ZSBvbjogNi85LzIwMTIgODowNDowNyBQTSBFU1QuPGJyIC8+DQo8L2Rpdj48L2Jv ZHk+PC9odG1sPg== ----boundary_107_ffa6a9ea-01dc-40f5-a50c-4c3b3d113f08--

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  • cf3 Can't stat ... in files.copyfrom promise

    - by Xerxes
    On the client: # cf-agent -KIv ... cf3 -> Handling file existence constraints on /etc/cfengine3 cf3 -> Copy file /etc/cfengine3 from /srv/cfengine/sysconf/server/inputs check cf3 No existing connection to 172.31.69.83 is established... cf3 Set cfengine port number to 5308 = 5308 cf3 -> Connect to 172.31.69.83 = 172.31.69.83 on port 5308 cf3 LastSaw host 172.31.69.83 now cf3 Loaded /var/lib/cfengine3/ppkeys/root-172.31.69.83.pub cf3 .....................[.h.a.i.l.]................................. cf3 Strong authentication of server=172.31.69.83 connection confirmed cf3 Server returned error: Unspecified server refusal (see verbose server output) cf3 Can't stat /srv/cfengine/sysconf/server/inputs in files.copyfrom promise cf3 ?> defining promise result class Cfengine_Inputs_Updated_Failed .... cf3 ......................................................... cf3 Promise handle: cf3 Promise made by: [cf-agent.cf ] FAILED 172.31.69.83:///srv/cfengine/sysconf/server/inputs -> localhost:///etc/cfengine3 However, on the server (172.31.69.83), there's no reason why it can't stat the directory: cyrus:/srv/cfengine/sysconf/server# ls -l /srv/cfengine/sysconf/server/inputs total 52 -rw-r--r-- 1 root root 2142 Sep 6 21:54 cf-agent.cf -rw-r--r-- 1 root root 831 Sep 6 18:31 cf-execd.cf -rw-r--r-- 1 root root 4517 Sep 6 21:44 cf-serverd.cf -rw-r--r-- 1 root root 3082 Sep 6 21:44 dns.cf -rw-r--r-- 1 root root 2028 Sep 6 15:12 failsafe.cf -rw-r--r-- 1 root root 5966 Sep 6 21:44 ldap-masters.cf -rw-r--r-- 1 root root 4380 Sep 6 18:31 ldap-security.cf -rw-r--r-- 1 root root 2735 Sep 6 08:21 lib-core.cf -rw-r--r-- 1 root root 1506 Sep 6 21:45 lib-utils.cf -rw-r--r-- 1 root root 2635 Sep 6 20:27 lib-vars.cf -rw-r--r-- 1 root root 2057 Sep 3 17:46 nss.cf -rw-r--r-- 1 root root 1472 Sep 6 18:31 packages.cf -rw-r--r-- 1 root root 1257 Sep 6 18:01 pam-security.cf -rw-r--r-- 1 root root 4019 Sep 6 19:32 promises.cf -rw-r--r-- 1 root root 2808 Sep 3 17:22 site.cf -rw-r--r-- 1 root root 1670 Sep 6 18:31 sudo-security.cf -rw-r--r-- 1 root root 831 Sep 6 18:31 sys-security.cf -rw-r--r-- 1 root root 890 Sep 6 18:31 sys-users.cf cyrus:/srv/cfengine/sysconf/server# I don't see anything interesting server side either when running: /usr/sbin/cf-serverd -d4 --verbose --no-fork And the following does not have any complaints: /usr/sbin/cf-promises -v Any ideas? I'm running cfengine3 on debian, v3.0.5+dfsg-1 - and the cf-agent.cf file is as follows: bundle agent Update { files: linux:: "${cf3.path[inputs]}" action => immediate, move_obstructions => "true", depth_search => Recursive, copy_from => MirrorFrom( "${cf3.host[server]}", "${cf3.path[scm-inputs]}", "true", "0400" ), classes => DefineSoftClass("Cfengine_Inputs_Updated") ; "${cf3.path[sbin]}" comment => "Setting cf3 client sbin scripts: ${cf3.path[sbin]}/", action => immediate, depth_search => Recursive, copy_from => MirrorFrom( "${cf3.host[server]}", "${cf3.path[scm-cnt-scripts]}", "false", "0555" ) ; reports: Cfengine_Inputs_Updated:: "[cf-agent.cf ] Services:CFAgent:Inputs:Updated"; Cfengine_Inputs_Updated_Failed:: "[cf-agent.cf ] FAILED ${cf3.host[server]}://${cf3.path[scm-inputs]} -> localhost://${cf3.path[inputs]}"; } I lie, there is something interesting with a little more debugging... AccessControl(/srv/cfengine/sysconf/server/inputs) AccessControl, match(/srv/cfengine/sysconf/server/inputs,client.com.au) encrypt request=1 Examining rule in access list (/srv/cfengine/sysconf/server/inputs,/home/cfengine)? cf3 Host client.com.au denied access to /srv/cfengine/sysconf/server/inputs Unappending Host client.com.au denied access to /srv/cfengine/sysconf/server/inputs cf3 Access control in sync Unappending Access control in sync Transaction Send[t 59][Packed text] Attempting to send 67 bytes SendSocketStream, sent 67 cf3 From (host=client.com.au,user=root,ip=172.31.69.3) Unappending From (host=client.com.au,user=root,ip=172.31.69.3) cf3 REFUSAL of request from connecting host: (SYNCH 1283777156 STAT /srv/cfengine/sysconf/server/inputs) Unappending REFUSAL of request from connecting host: (SYNCH 1283777156 STAT /srv/cfengine/sysconf/server/inputs) RecvSocketStream(8) cf3 -> Accepting a connection I'll keep looking.

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  • PPPTP VPN from Ubuntu cannot connect

    - by Andrea Polci
    I'm trying to configure under Linux (Kubuntu 9.10) a VPN I already use from Windows. I installed the network-manager-pptp package and added the vpn under Network Manager. These are the parameter under "advanced" button: Authentication Methods: PAP, CHAP, MSCHAP, SMCHAP2, EAP (I tried also with MSCHAP and MSCHAP2 only) Use MPPE Encryption: yes Crypto: Any Use stateful encryption: no Compression: Allow BSD compression: yes Allow Deflate compression: yes Allow TCP header compression: yes Send PPP echo packets: no When I try to connnect it doesn't work and this is what I get in the system log: 2010-04-08 13:53:47 pcelena NetworkManager <info> Starting VPN service 'org.freedesktop.NetworkManager.pptp'... 2010-04-08 13:53:47 pcelena NetworkManager <info> VPN service 'org.freedesktop.NetworkManager.pptp' started (org.freedesktop.NetworkManager.pptp), PID 4931 2010-04-08 13:53:47 pcelena NetworkManager <info> VPN service 'org.freedesktop.NetworkManager.pptp' just appeared, activating connections 2010-04-08 13:53:47 pcelena pppd[4932] Plugin /usr/lib/pppd/2.4.5//nm-pptp-pppd-plugin.so loaded. 2010-04-08 13:53:47 pcelena NetworkManager <info> VPN plugin state changed: 3 2010-04-08 13:53:47 pcelena pppd[4932] pppd 2.4.5 started by root, uid 0 2010-04-08 13:53:47 pcelena NetworkManager <info> VPN connection 'MYVPN' (Connect) reply received. 2010-04-08 13:53:47 pcelena NetworkManager SCPlugin-Ifupdown: devices added (path: /sys/devices/virtual/net/ppp0, iface: ppp0) 2010-04-08 13:53:47 pcelena NetworkManager SCPlugin-Ifupdown: device added (path: /sys/devices/virtual/net/ppp0, iface: ppp0): no ifupdown configuration found. 2010-04-08 13:53:47 pcelena pppd[4932] Using interface ppp0 2010-04-08 13:53:47 pcelena pppd[4932] Connect: ppp0 <--> /dev/pts/2 2010-04-08 13:53:47 pcelena pptp[4934] nm-pptp-service-4931 log[main:pptp.c:314]: The synchronous pptp option is NOT activated 2010-04-08 13:53:47 pcelena pptp[4927] nm-pptp-service-4918 log[ctrlp_rep:pptp_ctrl.c:251]: Sent control packet type is 7 'Outgoing-Call-Request' 2010-04-08 13:53:47 pcelena pptp[4927] nm-pptp-service-4918 log[ctrlp_disp:pptp_ctrl.c:858]: Received Outgoing Call Reply. 2010-04-08 13:53:47 pcelena pptp[4927] nm-pptp-service-4918 log[ctrlp_disp:pptp_ctrl.c:897]: Outgoing call established (call ID 1, peer's call ID 14800). 2010-04-08 13:53:48 pcelena pppd[4932] CHAP authentication succeeded 2010-04-08 13:53:48 pcelena pppd[4932] CHAP authentication succeeded 2010-04-08 13:53:48 pcelena pppd[4932] LCP terminated by peer 2010-04-08 13:53:48 pcelena pptp[4927] nm-pptp-service-4918 log[ctrlp_disp:pptp_ctrl.c:929]: Call disconnect notification received (call id 14800) 2010-04-08 13:53:48 pcelena pptp[4927] nm-pptp-service-4918 log[ctrlp_disp:pptp_ctrl.c:788]: Received Stop Control Connection Request. 2010-04-08 13:53:48 pcelena pptp[4927] nm-pptp-service-4918 log[ctrlp_rep:pptp_ctrl.c:251]: Sent control packet type is 4 'Stop-Control-Connection-Reply' 2010-04-08 13:53:48 pcelena pptp[4927] nm-pptp-service-4918 log[callmgr_main:pptp_callmgr.c:258]: Closing connection (shutdown) 2010-04-08 13:53:48 pcelena pptp[4927] nm-pptp-service-4918 log[ctrlp_rep:pptp_ctrl.c:251]: Sent control packet type is 12 'Call-Clear-Request' 2010-04-08 13:53:48 pcelena pptp[4927] nm-pptp-service-4918 log[callmgr_main:pptp_callmgr.c:258]: Closing connection (shutdown) 2010-04-08 13:53:48 pcelena pptp[4927] nm-pptp-service-4918 log[ctrlp_rep:pptp_ctrl.c:251]: Sent control packet type is 12 'Call-Clear-Request' 2010-04-08 13:53:48 pcelena pptp[4927] nm-pptp-service-4918 log[call_callback:pptp_callmgr.c:79]: Closing connection (call state) 2010-04-08 13:53:48 pcelena pppd[4932] Modem hangup 2010-04-08 13:53:48 pcelena pppd[4932] Connection terminated. 2010-04-08 13:53:48 pcelena NetworkManager <info> VPN plugin failed: 1 2010-04-08 13:53:48 pcelena NetworkManager SCPlugin-Ifupdown: devices removed (path: /sys/devices/virtual/net/ppp0, iface: ppp0) 2010-04-08 13:53:48 pcelena pppd[4932] Exit. 2010-04-08 13:53:48 pcelena NetworkManager <info> VPN plugin failed: 1 2010-04-08 13:53:48 pcelena NetworkManager <info> VPN plugin state changed: 6 2010-04-08 13:53:48 pcelena NetworkManager <info> VPN plugin state change reason: 0 2010-04-08 13:53:48 pcelena NetworkManager <WARN> connection_state_changed(): Could not process the request because no VPN connection was active. 2010-04-08 13:53:48 pcelena NetworkManager <info> Policy set 'Auto eth0' (eth0) as default for routing and DNS. 2010-04-08 13:54:01 pcelena NetworkManager <debug> [1270727641.001390] ensure_killed(): waiting for vpn service pid 4931 to exit 2010-04-08 13:54:01 pcelena NetworkManager <debug> [1270727641.001479] ensure_killed(): vpn service pid 4931 cleaned up Does anyone has suggestion on what can be the problem and how to make it work?

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