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  • SQL Server 2008 R2 Reporting Services - The Word is But a Stage (T-SQL Tuesday #006)

    - by smisner
    Host Michael Coles (blog|twitter) has selected LOB data as the topic for this month's T-SQL Tuesday, so I'll take this opportunity to post an overview of reporting with spatial data types. As part of my work with SQL Server 2008 R2 Reporting Services, I've been exploring the use of spatial data types in the new map data region. You can create a map using any of the following data sources: Map Gallery - a set of Shapefiles for the United States only that ships with Reporting Services ESRI Shapefile - a .shp file conforming to the Environmental Systems Research Institute, Inc. (ESRI) shapefile spatial data format SQL Server spatial data - a query that includes SQLGeography or SQLGeometry data types Rob Farley (blog|twitter) points out today in his T-SQL Tuesday post that using the SQL geography field is a preferable alternative to ESRI shapefiles for storing spatial data in SQL Server. So how do you get spatial data? If you don't already have a GIS application in-house, you can find a variety of sources. Here are a few to get you started: US Census Bureau Website, http://www.census.gov/geo/www/tiger/ Global Administrative Areas Spatial Database, http://biogeo.berkeley.edu/gadm/ Digital Chart of the World Data Server, http://www.maproom.psu.edu/dcw/ In a recent post by Pinal Dave (blog|twitter), you can find a link to free shapefiles for download and a tutorial for using Shape2SQL, a free tool to convert shapefiles into SQL Server data. In my post today, I'll show you how to use combine spatial data that describes boundaries with spatial data in AdventureWorks2008R2 that identifies stores locations to embed a map in a report. Preparing the spatial data First, I downloaded Shapefile data for the administrative boundaries in France and unzipped the data to a local folder. Then I used Shape2SQL to upload the data into a SQL Server database called Spatial. I'm not sure of the reason why, but I had to uncheck the option to create a spatial index to upload the data. Otherwise, the upload appeared to run successfully, but no table appeared in my database. The zip file that I downloaded contained three files, but I didn't know what was in them until I used Shape2SQL to upload the data into tables. Then I found that FRA_adm0 contains spatial data for the country of France, FRA_adm1 contains spatial data for each region, and FRA_adm2 contains spatial data for each department (a subdivision of region). Next I prepared my SQL query containing sales data for fictional stores selling Adventure Works products in France. The Person.Address table in the AdventureWorks2008R2 database (which you can download from Codeplex) contains a SpatialLocation column which I joined - along with several other tables - to the Sales.Customer and Sales.Store tables. I'll be able to superimpose this data on a map to see where these stores are located. I included the SQL script for this query (as well as the spatial data for France) in the downloadable project that I created for this post. Step 1: Using the Map Wizard to Create a Map of France You can build a map without using the wizard, but I find it's rather useful in this case. Whether you use Business Intelligence Development Studio (BIDS) or Report Builder 3.0, the map wizard is the same. I used BIDS so that I could create a project that includes all the files related to this post. To get started, I added an empty report template to the project and named it France Stores. Then I opened the Toolbox window and dragged the Map item to the report body which starts the wizard. Here are the steps to perform to create a map of France: On the Choose a source of spatial data page of the wizard, select SQL Server spatial query, and click Next. On the Choose a dataset with SQL Server spatial data page, select Add a new dataset with SQL Server spatial data. On the Choose a connection to a SQL Server spatial data source page, select New. In the Data Source Properties dialog box, on the General page, add a connecton string like this (changing your server name if necessary): Data Source=(local);Initial Catalog=Spatial Click OK and then click Next. On the Design a query page, add a query for the country shape, like this: select * from fra_adm1 Click Next. The map wizard reads the spatial data and renders it for you on the Choose spatial data and map view options page, as shown below. You have the option to add a Bing Maps layer which shows surrounding countries. Depending on the type of Bing Maps layer that you choose to add (from Road, Aerial, or Hybrid) and the zoom percentage you select, you can view city names and roads and various boundaries. To keep from cluttering my map, I'm going to omit the Bing Maps layer in this example, but I do recommend that you experiment with this feature. It's a nice integration feature. Use the + or - button to rexize the map as needed. (I used the + button to increase the size of the map until its edges were just inside the boundaries of the visible map area (which is called the viewport). You can eliminate the color scale and distance scale boxes that appear in the map area later. Select the Embed map data in this report for faster rendering. The spatial data won't be changing, so there's no need to leave it in the database. However, it does increase the size of the RDL. Click Next. On the Choose map visualization page, select Basic Map. We'll add data for visualization later. For now, we have just the outline of France to serve as the foundation layer for our map. Click Next, and then click Finish. Now click the color scale box in the lower left corner of the map, and press the Delete key to remove it. Then repeat to remove the distance scale box in the lower right corner of the map. Step 2: Add a Map Layer to an Existing Map The map data region allows you to add multiple layers. Each layer is associated with a different data set. Thus far, we have the spatial data that defines the regional boundaries in the first map layer. Now I'll add in another layer for the store locations by following these steps: If the Map Layers windows is not visible, click the report body, and then click twice anywhere on the map data region to display it. Click on the New Layer Wizard button in the Map layers window. And then we start over again with the process by choosing a spatial data source. Select SQL Server spatial query, and click Next. Select Add a new dataset with SQL Server spatial data, and click Next. Click New, add a connection string to the AdventureWorks2008R2 database, and click Next. Add a query with spatial data (like the one I included in the downloadable project), and click Next. The location data now appears as another layer on top of the regional map created earlier. Use the + button to resize the map again to fill as much of the viewport as possible without cutting off edges of the map. You might need to drag the map within the viewport to center it properly. Select Embed map data in this report, and click Next. On the Choose map visualization page, select Basic Marker Map, and click Next. On the Choose color theme and data visualization page, in the Marker drop-down list, change the marker to diamond. There's no particular reason for a diamond; I think it stands out a little better than a circle on this map. Clear the Single color map checkbox as another way to distinguish the markers from the map. You can of course create an analytical map instead, which would change the size and/or color of the markers according to criteria that you specify, such as sales volume of each store, but I'll save that exploration for another post on another day. Click Finish and then click Preview to see the rendered report. Et voilà...c'est fini. Yes, it's a very simple map at this point, but there are many other things you can do to enhance the map. I'll create a series of posts to explore the possibilities. Share this post: email it! | bookmark it! | digg it! | reddit! | kick it! | live it!

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  • SQL SERVER – How to Recover SQL Database Data Deleted by Accident

    - by Pinal Dave
    In Repair a SQL Server database using a transaction log explorer, I showed how to use ApexSQL Log, a SQL Server transaction log viewer, to recover a SQL Server database after a disaster. In this blog, I’ll show you how to use another SQL Server disaster recovery tool from ApexSQL in a situation when data is accidentally deleted. You can download ApexSQL Recover here, install, and play along. With a good SQL Server disaster recovery strategy, data recovery is not a problem. You have a reliable full database backup with valid data, a full database backup and subsequent differential database backups, or a full database backup and a chain of transaction log backups. But not all situations are ideal. Here we’ll address some sub-optimal scenarios, where you can still successfully recover data. If you have only a full database backup This is the least optimal SQL Server disaster recovery strategy, as it doesn’t ensure minimal data loss. For example, data was deleted on Wednesday. Your last full database backup was created on Sunday, three days before the records were deleted. By using the full database backup created on Sunday, you will be able to recover SQL database records that existed in the table on Sunday. If there were any records inserted into the table on Monday or Tuesday, they will be lost forever. The same goes for records modified in this period. This method will not bring back modified records, only the old records that existed on Sunday. If you restore this full database backup, all your changes (intentional and accidental) will be lost and the database will be reverted to the state it had on Sunday. What you have to do is compare the records that were in the table on Sunday to the records on Wednesday, create a synchronization script, and execute it against the Wednesday database. If you have a full database backup followed by differential database backups Let’s say the situation is the same as in the example above, only you create a differential database backup every night. Use the full database backup created on Sunday, and the last differential database backup (created on Tuesday). In this scenario, you will lose only the data inserted and updated after the differential backup created on Tuesday. If you have a full database backup and a chain of transaction log backups This is the SQL Server disaster recovery strategy that provides minimal data loss. With a full chain of transaction logs, you can recover the SQL database to an exact point in time. To provide optimal results, you have to know exactly when the records were deleted, because restoring to a later point will not bring back the records. This method requires restoring the full database backup first. If you have any differential log backup created after the last full database backup, restore the most recent one. Then, restore transaction log backups, one by one, it the order they were created starting with the first created after the restored differential database backup. Now, the table will be in the state before the records were deleted. You have to identify the deleted records, script them and run the script against the original database. Although this method is reliable, it is time-consuming and requires a lot of space on disk. How to easily recover deleted records? The following solution enables you to recover SQL database records even if you have no full or differential database backups and no transaction log backups. To understand how ApexSQL Recover works, I’ll explain what happens when table data is deleted. Table data is stored in data pages. When you delete table records, they are not immediately deleted from the data pages, but marked to be overwritten by new records. Such records are not shown as existing anymore, but ApexSQL Recover can read them and create undo script for them. How long will deleted records stay in the MDF file? It depends on many factors, as time passes it’s less likely that the records will not be overwritten. The more transactions occur after the deletion, the more chances the records will be overwritten and permanently lost. Therefore, it’s recommended to create a copy of the database MDF and LDF files immediately (if you cannot take your database offline until the issue is solved) and run ApexSQL Recover on them. Note that a full database backup will not help here, as the records marked for overwriting are not included in the backup. First, I’ll delete some records from the Person.EmailAddress table in the AdventureWorks database.   I can delete these records in SQL Server Management Studio, or execute a script such as DELETE FROM Person.EmailAddress WHERE BusinessEntityID BETWEEN 70 AND 80 Then, I’ll start ApexSQL Recover and select From DELETE operation in the Recovery tab.   In the Select the database to recover step, first select the SQL Server instance. If it’s not shown in the drop-down list, click the Server icon right to the Server drop-down list and browse for the SQL Server instance, or type the instance name manually. Specify the authentication type and select the database in the Database drop-down list.   In the next step, you’re prompted to add additional data sources. As this can be a tricky step, especially for new users, ApexSQL Recover offers help via the Help me decide option.   The Help me decide option guides you through a series of questions about the database transaction log and advises what files to add. If you know that you have no transaction log backups or detached transaction logs, or the online transaction log file has been truncated after the data was deleted, select No additional transaction logs are available. If you know that you have transaction log backups that contain the delete transactions you want to recover, click Add transaction logs. The online transaction log is listed and selected automatically.   Click Add if to add transaction log backups. It would be best if you have a full transaction log chain, as explained above. The next step for this option is to specify the time range.   Selecting a small time range for the time of deletion will create the recovery script just for the accidentally deleted records. A wide time range might script the records deleted on purpose, and you don’t want that. If needed, you can check the script generated and manually remove such records. After that, for all data sources options, the next step is to select the tables. Be careful here, if you deleted some data from other tables on purpose, and don’t want to recover them, don’t select all tables, as ApexSQL Recover will create the INSERT script for them too.   The next step offers two options: to create a recovery script that will insert the deleted records back into the Person.EmailAddress table, or to create a new database, create the Person.EmailAddress table in it, and insert the deleted records. I’ll select the first one.   The recovery process is completed and 11 records are found and scripted, as expected.   To see the script, click View script. ApexSQL Recover has its own script editor, where you can review, modify, and execute the recovery script. The insert into statements look like: INSERT INTO Person.EmailAddress( BusinessEntityID, EmailAddressID, EmailAddress, rowguid, ModifiedDate) VALUES( 70, 70, N'[email protected]' COLLATE SQL_Latin1_General_CP1_CI_AS, 'd62c5b4e-c91f-403f-b630-7b7e0fda70ce', '20030109 00:00:00.000' ); To execute the script, click Execute in the menu.   If you want to check whether the records are really back, execute SELECT * FROM Person.EmailAddress WHERE BusinessEntityID BETWEEN 70 AND 80 As shown, ApexSQL Recover recovers SQL database data after accidental deletes even without the database backup that contains the deleted data and relevant transaction log backups. ApexSQL Recover reads the deleted data from the database data file, so this method can be used even for databases in the Simple recovery model. Besides recovering SQL database records from a DELETE statement, ApexSQL Recover can help when the records are lost due to a DROP TABLE, or TRUNCATE statement, as well as repair a corrupted MDF file that cannot be attached to as SQL Server instance. You can find more information about how to recover SQL database lost data and repair a SQL Server database on ApexSQL Solution center. There are solutions for various situations when data needs to be recovered. Reference: Pinal Dave (http://blog.sqlauthority.com)Filed under: PostADay, SQL, SQL Authority, SQL Backup and Restore, SQL Query, SQL Server, SQL Tips and Tricks, T SQL

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  • Use an Ubuntu Live CD to Securely Wipe Your PC’s Hard Drive

    - by Trevor Bekolay
    Deleting files or quickly formatting a drive isn’t enough for sensitive personal information. We’ll show you how to get rid of it for good using a Ubuntu Live CD. When you delete a file in Windows, Ubuntu, or any other operating system, it doesn’t actually destroy the data stored on your hard drive, it just marks that data as “deleted.” If you overwrite it later, then that data is generally unrecoverable, but if the operating system don’t happen to overwrite it, then your data is still stored on your hard drive, recoverable by anyone who has the right software. By securely delete files or entire hard drives, your data will be gone for good. Note: Modern hard drives are extremely sophisticated, as are the experts who recover data for a living. There is no guarantee that the methods covered in this article will make your data completely unrecoverable; however, they will make your data unrecoverable to the majority of recovery methods, and all methods that are readily available to the general public. Shred individual files Most of the data stored on your hard drive is harmless, and doesn’t reveal anything about you. If there are just a few files that you know you don’t want someone else to see, then the easiest way to get rid of them is a built-in Linux utility called shred. Open a terminal window by clicking on Applications at the top-left of the screen, then expanding the Accessories menu and clicking on Terminal. Navigate to the file that you want to delete using cd to change directories and ls to list the files and folders in the current directory. As an example, we’ve got a file called BankInfo.txt on a Windows NTFS-formatted hard drive. We want to delete it securely, so we’ll call shred by entering the following in the terminal window: shred <file> which is, in our example: shred BankInfo.txt Notice that our BankInfo.txt file still exists, even though we’ve shredded it. A quick look at the contents of BankInfo.txt make it obvious that the file has indeed been securely overwritten. We can use some command-line arguments to make shred delete the file from the hard drive as well. We can also be extra-careful about the shredding process by upping the number of times shred overwrites the original file. To do this, in the terminal, type in: shred –remove –iterations=<num> <file> By default, shred overwrites the file 25 times. We’ll double this, giving us the following command: shred –remove –iterations=50 BankInfo.txt BankInfo.txt has now been securely wiped on the physical disk, and also no longer shows up in the directory listing. Repeat this process for any sensitive files on your hard drive! Wipe entire hard drives If you’re disposing of an old hard drive, or giving it to someone else, then you might instead want to wipe your entire hard drive. shred can be invoked on hard drives, but on modern file systems, the shred process may be reversible. We’ll use the program wipe to securely delete all of the data on a hard drive. Unlike shred, wipe is not included in Ubuntu by default, so we have to install it. Open up the Synaptic Package Manager by clicking on System in the top-left corner of the screen, then expanding the Administration folder and clicking on Synaptic Package Manager. wipe is part of the Universe repository, which is not enabled by default. We’ll enable it by clicking on Settings > Repositories in the Synaptic Package Manager window. Check the checkbox next to “Community-maintained Open Source software (universe)”. Click Close. You’ll need to reload Synaptic’s package list. Click on the Reload button in the main Synaptic Package Manager window. Once the package list has been reloaded, the text over the search field will change to “Rebuilding search index”. Wait until it reads “Quick search,” and then type “wipe” into the search field. The wipe package should come up, along with some other packages that perform similar functions. Click on the checkbox to the left of the label “wipe” and select “Mark for Installation”. Click on the Apply button to start the installation process. Click the Apply button on the Summary window that pops up. Once the installation is done, click the Close button and close the Synaptic Package Manager window. Open a terminal window by clicking on Applications in the top-left of the screen, then Accessories > Terminal. You need to figure our the correct hard drive to wipe. If you wipe the wrong hard drive, that data will not be recoverable, so exercise caution! In the terminal window, type in: sudo fdisk -l A list of your hard drives will show up. A few factors will help you identify the right hard drive. One is the file system, found in the System column of  the list – Windows hard drives are usually formatted as NTFS (which shows up as HPFS/NTFS). Another good identifier is the size of the hard drive, which appears after its identifier (highlighted in the following screenshot). In our case, the hard drive we want to wipe is only around 1 GB large, and is formatted as NTFS. We make a note of the label found under the the Device column heading. If you have multiple partitions on this hard drive, then there will be more than one device in this list. The wipe developers recommend wiping each partition separately. To start the wiping process, type the following into the terminal: sudo wipe <device label> In our case, this is: sudo wipe /dev/sda1 Again, exercise caution – this is the point of no return! Your hard drive will be completely wiped. It may take some time to complete, depending on the size of the drive you’re wiping. Conclusion If you have sensitive information on your hard drive – and chances are you probably do – then it’s a good idea to securely delete sensitive files before you give away or dispose of your hard drive. The most secure way to delete your data is with a few swings of a hammer, but shred and wipe from a Ubuntu Live CD is a good alternative! Similar Articles Productive Geek Tips Reset Your Ubuntu Password Easily from the Live CDScan a Windows PC for Viruses from a Ubuntu Live CDRecover Deleted Files on an NTFS Hard Drive from a Ubuntu Live CDCreate a Bootable Ubuntu 9.10 USB Flash DriveCreate a Bootable Ubuntu USB Flash Drive the Easy Way TouchFreeze Alternative in AutoHotkey The Icy Undertow Desktop Windows Home Server – Backup to LAN The Clear & Clean Desktop Use This Bookmarklet to Easily Get Albums Use AutoHotkey to Assign a Hotkey to a Specific Window Latest Software Reviews Tinyhacker Random Tips DVDFab 6 Revo Uninstaller Pro Registry Mechanic 9 for Windows PC Tools Internet Security Suite 2010 Office 2010 Product Guides Google Maps Place marks – Pizza, Guns or Strip Clubs Monitor Applications With Kiwi LocPDF is a Visual PDF Search Tool Download Free iPad Wallpapers at iPad Decor Get Your Delicious Bookmarks In Firefox’s Awesome Bar

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  • Change or Reset Windows Password from a Ubuntu Live CD

    - by Trevor Bekolay
    If you can’t log in even after trying your twelve passwords, or you’ve inherited a computer complete with password-protected profiles, worry not – you don’t have to do a fresh install of Windows. We’ll show you how to change or reset your Windows password from a Ubuntu Live CD. This method works for all of the NT-based version of Windows – anything from Windows 2000 and later, basically. And yes, that includes Windows 7. You’ll need a Ubuntu 9.10 Live CD, or a bootable Ubuntu 9.10 Flash Drive. If you don’t have one, or have forgotten how to boot from the flash drive, check out our article on creating a bootable Ubuntu 9.10 flash drive. The program that lets us manipulate Windows passwords is called chntpw. The steps to install it are different in 32-bit and 64-bit versions of Ubuntu. Installation: 32-bit Open up Synaptic Package Manager by clicking on System at the top of the screen, expanding the Administration section, and clicking on Synaptic Package Manager. chntpw is found in the universe repository. Repositories are a way for Ubuntu to group software together so that users are able to choose if they want to use only completely open source software maintained by Ubuntu developers, or branch out and use software with different licenses and maintainers. To enable software from the universe repository, click on Settings > Repositories in the Synaptic window. Add a checkmark beside the box labeled “Community-maintained Open Source software (universe)” and then click close. When you change the repositories you are selecting software from, you have to reload the list of available software. In the main Synaptic window, click on the Reload button. The software lists will be downloaded. Once downloaded, Synaptic must rebuild its search index. The label over the text field by the Search button will read “Rebuilding search index.” When it reads “Quick search,” type chntpw in the text field. The package will show up in the list. Click on the checkbox near the chntpw name. Click on Mark for Installation. chntpw won’t actually be installed until you apply the changes you’ve made, so click on the Apply button in the Synaptic window now. You will be prompted to accept the changes. Click Apply. The changes should be applied quickly. When they’re done, click Close. chntpw is now installed! You can close Synaptic Package Manager. Skip to the section titled Using chntpw to reset your password. Installation: 64-bit The version of chntpw available in Ubuntu’s universe repository will not work properly on a 64-bit machine. Fortunately, a patched version exists in Debian’s Unstable branch, so let’s download it from there and install it manually. Open Firefox. Whether it’s your preferred browser or not, it’s very readily accessible in the Ubuntu Live CD environment, so it will be the easiest to use. There’s a shortcut to Firefox in the top panel. Navigate to http://packages.debian.org/sid/amd64/chntpw/download and download the latest version of chntpw for 64-bit machines. Note: In most cases it would be best to add the Debian Unstable branch to a package manager, but since the Live CD environment will revert to its original state once you reboot, it’ll be faster to just download the .deb file. Save the .deb file to the default location. You can close Firefox if desired. Open a terminal window by clicking on Applications at the top-left of the screen, expanding the Accessories folder, and clicking on Terminal. In the terminal window, enter the following text, hitting enter after each line: cd Downloadssudo dpkg –i chntpw* chntpw will now be installed. Using chntpw to reset your password Before running chntpw, you will have to mount the hard drive that contains your Windows installation. In most cases, Ubuntu 9.10 makes this simple. Click on Places at the top-left of the screen. If your Windows drive is easily identifiable – usually by its size – then left click on it. If it is not obvious, then click on Computer and check out each hard drive until you find the correct one. The correct hard drive will have the WINDOWS folder in it. When you find it, make a note of the drive’s label that appears in the menu bar of the file browser. If you don’t already have one open, start a terminal window by going to Applications > Accessories > Terminal. In the terminal window, enter the commands cd /medials pressing enter after each line. You should see one or more strings of text appear; one of those strings should correspond with the string that appeared in the title bar of the file browser earlier. Change to that directory by entering the command cd <hard drive label> Since the hard drive label will be very annoying to type in, you can use a shortcut by typing in the first few letters or numbers of the drive label (capitalization matters) and pressing the Tab key. It will automatically complete the rest of the string (if those first few letters or numbers are unique). We want to switch to a certain Windows directory. Enter the command: cd WINDOWS/system32/config/ Again, you can use tab-completion to speed up entering this command. To change or reset the administrator password, enter: sudo chntpw SAM SAM is the file that contains your Windows registry. You will see some text appear, including a list of all of the users on your system. At the bottom of the terminal window, you should see a prompt that begins with “User Edit Menu:” and offers four choices. We recommend that you clear the password to blank (you can always set a new password in Windows once you log in). To do this, enter “1” and then “y” to confirm. If you would like to change the password instead, enter “2”, then your desired password, and finally “y” to confirm. If you would like to reset or change the password of a user other than the administrator, enter: sudo chntpw –u <username> SAM From here, you can follow the same steps as before: enter “1” to reset the password to blank, or “2” to change it to a value you provide. And that’s it! Conclusion chntpw is a very useful utility provided for free by the open source community. It may make you think twice about how secure the Windows login system is, but knowing how to use chntpw can save your tail if your memory fails you two or eight times! Similar Articles Productive Geek Tips Reset Your Ubuntu Password Easily from the Live CDChange Your Forgotten Windows Password with the Linux System Rescue CDHow to Create and Use a Password Reset Disk in Windows Vista & Windows 7Reset Your Forgotten Password the Easy Way Using the Ultimate Boot CD for WindowsHow to install Spotify in Ubuntu 9.10 using Wine TouchFreeze Alternative in AutoHotkey The Icy Undertow Desktop Windows Home Server – Backup to LAN The Clear & Clean Desktop Use This Bookmarklet to Easily Get Albums Use AutoHotkey to Assign a Hotkey to a Specific Window Latest Software Reviews Tinyhacker Random Tips DVDFab 6 Revo Uninstaller Pro Registry Mechanic 9 for Windows PC Tools Internet Security Suite 2010 Add a Custom Title in IE using Spybot or Spyware Blaster When You Need to Hail a Taxi in NYC Live Map of Marine Traffic NoSquint Remembers Site Specific Zoom Levels (Firefox) New Firefox release 3.6.3 fixes 1 Critical bug Dark Side of the Moon (8-bit)

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  • Recover Deleted Files on an NTFS Hard Drive from a Ubuntu Live CD

    - by Trevor Bekolay
    Accidentally deleting a file is a terrible feeling. Not being able to boot into Windows and undelete that file makes that even worse. Fortunately, you can recover deleted files on NTFS hard drives from an Ubuntu Live CD. To show this process, we created four files on the desktop of a Windows XP machine, and then deleted them. We then booted up the same machine with the bootable Ubuntu 9.10 USB Flash Drive that we created last week. Once Ubuntu 9.10 boots up, open a terminal by clicking Applications in the top left of the screen, and then selecting Accessories > Terminal. To undelete our files, we first need to identify the hard drive that we want to undelete from. In the terminal window, type in: sudo fdisk –l and press enter. What you’re looking for is a line that ends with HPSF/NTFS (under the heading System). In our case, the device is “/dev/sda1”. This may be slightly different for you, but it will still begin with /dev/. Note this device name. If you have more than one hard drive partition formatted as NTFS, then you may be able to identify the correct partition by the size. If you look at the second line of text in the screenshot above, it reads “Disk /dev/sda: 136.4 GB, …” This means that the hard drive that Ubuntu has named /dev/sda is 136.4 GB large. If your hard drives are of different size, then this information can help you track down the right device name to use. Alternatively, you can just try them all, though this can be time consuming for large hard drives. Now that you know the name Ubuntu has assigned to your hard drive, we’ll scan it to see what files we can uncover. In the terminal window, type: sudo ntfsundelete <HD name> and hit enter. In our case, the command is: sudo ntfsundelete /dev/sda1 The names of files that can recovered show up in the far right column. The percentage in the third column tells us how much of that file can be recovered. Three of the four files that we originally deleted are showing up in this list, even though we shut down the computer right after deleting the four files – so even in ideal cases, your files may not be recoverable. Nevertheless, we have three files that we can recover – two JPGs and an MPG. Note: ntfsundelete is immediately available in the Ubuntu 9.10 Live CD. If you are in a different version of Ubuntu, or for some other reason get an error when trying to use ntfsundelete, you can install it by entering “sudo apt-get install ntfsprogs” in a terminal window. To quickly recover the two JPGs, we will use the * wildcard to recover all of the files that end with .jpg. In the terminal window, enter sudo ntfsundelete <HD name> –u –m *.jpg which is, in our case, sudo ntfsundelete /dev/sda1 –u –m *.jpg The two files are recovered from the NTFS hard drive and saved in the current working directory of the terminal. By default, this is the home directory of the current user, though we are working in the Desktop folder. Note that the ntfsundelete program does not make any changes to the original NTFS hard drive. If you want to take those files and put them back in the NTFS hard drive, you will have to move them there after they are undeleted with ntfsundelete. Of course, you can also put them on your flash drive or open Firefox and email them to yourself – the sky’s the limit! We have one more file to undelete – our MPG. Note the first column on the far left. It contains a number, its Inode. Think of this as the file’s unique identifier. Note this number. To undelete a file by its Inode, enter the following in the terminal: sudo ntfsundelete <HD name> –u –i <Inode> In our case, this is: sudo ntfsundelete /dev/sda1 –u –i 14159 This recovers the file, along with an identifier that we don’t really care about. All three of our recoverable files are now recovered. However, Ubuntu lets us know visually that we can’t use these files yet. That’s because the ntfsundelete program saves the files as the “root” user, not the “ubuntu” user. We can verify this by typing the following in our terminal window: ls –l We want these three files to be owned by ubuntu, not root. To do this, enter the following in the terminal window: sudo chown ubuntu <Files> If the current folder has other files in it, you may not want to change their owner to ubuntu. However, in our case, we only have these three files in this folder, so we will use the * wildcard to change the owner of all three files. sudo chown ubuntu * The files now look normal, and we can do whatever we want with them. Hopefully you won’t need to use this tip, but if you do, ntfsundelete is a nice command-line utility. It doesn’t have a fancy GUI like many of the similar Windows programs, but it is a powerful tool that can recover your files quickly. See ntfsundelete’s manual page for more detailed usage information Similar Articles Productive Geek Tips Reset Your Ubuntu Password Easily from the Live CDUse Ubuntu Live CD to Backup Files from Your Dead Windows ComputerCreate a Bootable Ubuntu 9.10 USB Flash DriveCreate a Bootable Ubuntu USB Flash Drive the Easy WayGuide to Using Check Disk in Windows Vista TouchFreeze Alternative in AutoHotkey The Icy Undertow Desktop Windows Home Server – Backup to LAN The Clear & Clean Desktop Use This Bookmarklet to Easily Get Albums Use AutoHotkey to Assign a Hotkey to a Specific Window Latest Software Reviews Tinyhacker Random Tips Revo Uninstaller Pro Registry Mechanic 9 for Windows PC Tools Internet Security Suite 2010 PCmover Professional Windows 7 Easter Theme YoWindoW, a real time weather screensaver Optimize your computer the Microsoft way Stormpulse provides slick, real time weather data Geek Parents – Did you try Parental Controls in Windows 7? Change DNS servers on the fly with DNS Jumper

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  • Is Berkeley DB a NoSQL solution?

    - by Gregory Burd
    Berkeley DB is a library. To use it to store data you must link the library into your application. You can use most programming languages to access the API, the calls across these APIs generally mimic the Berkeley DB C-API which makes perfect sense because Berkeley DB is written in C. The inspiration for Berkeley DB was the DBM library, a part of the earliest versions of UNIX written by AT&T's Ken Thompson in 1979. DBM was a simple key/value hashtable-based storage library. In the early 1990s as BSD UNIX was transitioning from version 4.3 to 4.4 and retrofitting commercial code owned by AT&T with unencumbered code, it was the future founders of Sleepycat Software who wrote libdb (aka Berkeley DB) as the replacement for DBM. The problem it addressed was fast, reliable local key/value storage. At that time databases almost always lived on a single node, even the most sophisticated databases only had simple fail-over two node solutions. If you had a lot of data to store you would choose between the few commercial RDBMS solutions or to write your own custom solution. Berkeley DB took the headache out of the custom approach. These basic market forces inspired other DBM implementations. There was the "New DBM" (ndbm) and the "GNU DBM" (GDBM) and a few others, but the theme was the same. Even today TokyoCabinet calls itself "a modern implementation of DBM" mimicking, and improving on, something first created over thirty years ago. In the mid-1990s, DBM was the name for what you needed if you were looking for fast, reliable local storage. Fast forward to today. What's changed? Systems are connected over fast, very reliable networks. Disks are cheep, fast, and capable of storing huge amounts of data. CPUs continued to follow Moore's Law, processing power that filled a room in 1990 now fits in your pocket. PCs, servers, and other computers proliferated both in business and the personal markets. In addition to the new hardware entire markets, social systems, and new modes of interpersonal communication moved onto the web and started evolving rapidly. These changes cause a massive explosion of data and a need to analyze and understand that data. Taken together this resulted in an entirely different landscape for database storage, new solutions were needed. A number of novel solutions stepped up and eventually a category called NoSQL emerged. The new market forces inspired the CAP theorem and the heated debate of BASE vs. ACID. But in essence this was simply the market looking at what to trade off to meet these new demands. These new database systems shared many qualities in common. There were designed to address massive amounts of data, millions of requests per second, and scale out across multiple systems. The first large-scale and successful solution was Dynamo, Amazon's distributed key/value database. Dynamo essentially took the next logical step and added a twist. Dynamo was to be the database of record, it would be distributed, data would be partitioned across many nodes, and it would tolerate failure by avoiding single points of failure. Amazon did this because they recognized that the majority of the dynamic content they provided to customers visiting their web store front didn't require the services of an RDBMS. The queries were simple, key/value look-ups or simple range queries with only a few queries that required more complex joins. They set about to use relational technology only in places where it was the best solution for the task, places like accounting and order fulfillment, but not in the myriad of other situations. The success of Dynamo, and it's design, inspired the next generation of Non-SQL, distributed database solutions including Cassandra, Riak and Voldemort. The problem their designers set out to solve was, "reliability at massive scale" so the first focal point was distributed database algorithms. Underneath Dynamo there is a local transactional database; either Berkeley DB, Berkeley DB Java Edition, MySQL or an in-memory key/value data structure. Dynamo was an evolution of local key/value storage onto networks. Cassandra, Riak, and Voldemort all faced similar design decisions and one, Voldemort, choose Berkeley DB Java Edition for it's node-local storage. Riak at first was entirely in-memory, but has recently added write-once, append-only log-based on-disk storage similar type of storage as Berkeley DB except that it is based on a hash table which must reside entirely in-memory rather than a btree which can live in-memory or on disk. Berkeley DB evolved too, we added high availability (HA) and a replication manager that makes it easy to setup replica groups. Berkeley DB's replication doesn't partitioned the data, every node keeps an entire copy of the database. For consistency, there is a single node where writes are committed first - a master - then those changes are delivered to the replica nodes as log records. Applications can choose to wait until all nodes are consistent, or fire and forget allowing Berkeley DB to eventually become consistent. Berkeley DB's HA scales-out quite well for read-intensive applications and also effectively eliminates the central point of failure by allowing replica nodes to be elected (using a PAXOS algorithm) to mastership if the master should fail. This implementation covers a wide variety of use cases. MemcacheDB is a server that implements the Memcache network protocol but uses Berkeley DB for storage and HA to replicate the cache state across all the nodes in the cache group. Google Accounts, the user authentication layer for all Google properties, was until recently running Berkeley DB HA. That scaled to a globally distributed system. That said, most NoSQL solutions try to partition (shard) data across nodes in the replication group and some allow writes as well as reads at any node, Berkeley DB HA does not. So, is Berkeley DB a "NoSQL" solution? Not really, but it certainly is a component of many of the existing NoSQL solutions out there. Forgetting all the noise about how NoSQL solutions are complex distributed databases when you boil them down to a single node you still have to store the data to some form of stable local storage. DBMs solved that problem a long time ago. NoSQL has more to do with the layers on top of the DBM; the distributed, sometimes-consistent, partitioned, scale-out storage that manage key/value or document sets and generally have some form of simple HTTP/REST-style network API. Does Berkeley DB do that? Not really. Is Berkeley DB a "NoSQL" solution today? Nope, but it's the most robust solution on which to build such a system. Re-inventing the node-local data storage isn't easy. A lot of people are starting to come to appreciate the sophisticated features found in Berkeley DB, even mimic them in some cases. Could Berkeley DB grow into a NoSQL solution? Absolutely. Our key/value API could be extended over the net using any of a number of existing network protocols such as memcache or HTTP/REST. We could adapt our node-local data partitioning out over replicated nodes. We even have a nice query language and cost-based query optimizer in our BDB XML product that we could reuse were we to build out a document-based NoSQL-style product. XML and JSON are not so different that we couldn't adapt one to work with the other interchangeably. Without too much effort we could add what's missing, we could jump into this No SQL market withing a single product development cycle. Why isn't Berkeley DB already a NoSQL solution? Why aren't we working on it? Why indeed...

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  • SQL SERVER – Guest Post – Architecting Data Warehouse – Niraj Bhatt

    - by pinaldave
    Niraj Bhatt works as an Enterprise Architect for a Fortune 500 company and has an innate passion for building / studying software systems. He is a top rated speaker at various technical forums including Tech·Ed, MCT Summit, Developer Summit, and Virtual Tech Days, among others. Having run a successful startup for four years Niraj enjoys working on – IT innovations that can impact an enterprise bottom line, streamlining IT budgets through IT consolidation, architecture and integration of systems, performance tuning, and review of enterprise applications. He has received Microsoft MVP award for ASP.NET, Connected Systems and most recently on Windows Azure. When he is away from his laptop, you will find him taking deep dives in automobiles, pottery, rafting, photography, cooking and financial statements though not necessarily in that order. He is also a manager/speaker at BDOTNET, Asia’s largest .NET user group. Here is the guest post by Niraj Bhatt. As data in your applications grows it’s the database that usually becomes a bottleneck. It’s hard to scale a relational DB and the preferred approach for large scale applications is to create separate databases for writes and reads. These databases are referred as transactional database and reporting database. Though there are tools / techniques which can allow you to create snapshot of your transactional database for reporting purpose, sometimes they don’t quite fit the reporting requirements of an enterprise. These requirements typically are data analytics, effective schema (for an Information worker to self-service herself), historical data, better performance (flat data, no joins) etc. This is where a need for data warehouse or an OLAP system arises. A Key point to remember is a data warehouse is mostly a relational database. It’s built on top of same concepts like Tables, Rows, Columns, Primary keys, Foreign Keys, etc. Before we talk about how data warehouses are typically structured let’s understand key components that can create a data flow between OLTP systems and OLAP systems. There are 3 major areas to it: a) OLTP system should be capable of tracking its changes as all these changes should go back to data warehouse for historical recording. For e.g. if an OLTP transaction moves a customer from silver to gold category, OLTP system needs to ensure that this change is tracked and send to data warehouse for reporting purpose. A report in context could be how many customers divided by geographies moved from sliver to gold category. In data warehouse terminology this process is called Change Data Capture. There are quite a few systems that leverage database triggers to move these changes to corresponding tracking tables. There are also out of box features provided by some databases e.g. SQL Server 2008 offers Change Data Capture and Change Tracking for addressing such requirements. b) After we make the OLTP system capable of tracking its changes we need to provision a batch process that can run periodically and takes these changes from OLTP system and dump them into data warehouse. There are many tools out there that can help you fill this gap – SQL Server Integration Services happens to be one of them. c) So we have an OLTP system that knows how to track its changes, we have jobs that run periodically to move these changes to warehouse. The question though remains is how warehouse will record these changes? This structural change in data warehouse arena is often covered under something called Slowly Changing Dimension (SCD). While we will talk about dimensions in a while, SCD can be applied to pure relational tables too. SCD enables a database structure to capture historical data. This would create multiple records for a given entity in relational database and data warehouses prefer having their own primary key, often known as surrogate key. As I mentioned a data warehouse is just a relational database but industry often attributes a specific schema style to data warehouses. These styles are Star Schema or Snowflake Schema. The motivation behind these styles is to create a flat database structure (as opposed to normalized one), which is easy to understand / use, easy to query and easy to slice / dice. Star schema is a database structure made up of dimensions and facts. Facts are generally the numbers (sales, quantity, etc.) that you want to slice and dice. Fact tables have these numbers and have references (foreign keys) to set of tables that provide context around those facts. E.g. if you have recorded 10,000 USD as sales that number would go in a sales fact table and could have foreign keys attached to it that refers to the sales agent responsible for sale and to time table which contains the dates between which that sale was made. These agent and time tables are called dimensions which provide context to the numbers stored in fact tables. This schema structure of fact being at center surrounded by dimensions is called Star schema. A similar structure with difference of dimension tables being normalized is called a Snowflake schema. This relational structure of facts and dimensions serves as an input for another analysis structure called Cube. Though physically Cube is a special structure supported by commercial databases like SQL Server Analysis Services, logically it’s a multidimensional structure where dimensions define the sides of cube and facts define the content. Facts are often called as Measures inside a cube. Dimensions often tend to form a hierarchy. E.g. Product may be broken into categories and categories in turn to individual items. Category and Items are often referred as Levels and their constituents as Members with their overall structure called as Hierarchy. Measures are rolled up as per dimensional hierarchy. These rolled up measures are called Aggregates. Now this may seem like an overwhelming vocabulary to deal with but don’t worry it will sink in as you start working with Cubes and others. Let’s see few other terms that we would run into while talking about data warehouses. ODS or an Operational Data Store is a frequently misused term. There would be few users in your organization that want to report on most current data and can’t afford to miss a single transaction for their report. Then there is another set of users that typically don’t care how current the data is. Mostly senior level executives who are interesting in trending, mining, forecasting, strategizing, etc. don’t care for that one specific transaction. This is where an ODS can come in handy. ODS can use the same star schema and the OLAP cubes we saw earlier. The only difference is that the data inside an ODS would be short lived, i.e. for few months and ODS would sync with OLTP system every few minutes. Data warehouse can periodically sync with ODS either daily or weekly depending on business drivers. Data marts are another frequently talked about topic in data warehousing. They are subject-specific data warehouse. Data warehouses that try to span over an enterprise are normally too big to scope, build, manage, track, etc. Hence they are often scaled down to something called Data mart that supports a specific segment of business like sales, marketing, or support. Data marts too, are often designed using star schema model discussed earlier. Industry is divided when it comes to use of data marts. Some experts prefer having data marts along with a central data warehouse. Data warehouse here acts as information staging and distribution hub with spokes being data marts connected via data feeds serving summarized data. Others eliminate the need for a centralized data warehouse citing that most users want to report on detailed data. Reference: Pinal Dave (http://blog.SQLAuthority.com) Filed under: Best Practices, Business Intelligence, Data Warehousing, Database, Pinal Dave, PostADay, Readers Contribution, SQL, SQL Authority, SQL Query, SQL Server, SQL Tips and Tricks, T SQL, Technology

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  • The Proper Use of the VM Role in Windows Azure

    - by BuckWoody
    At the Professional Developer’s Conference (PDC) in 2010 we announced an addition to the Computational Roles in Windows Azure, called the VM Role. This new feature allows a great deal of control over the applications you write, but some have confused it with our full infrastructure offering in Windows Hyper-V. There is a proper architecture pattern for both of them. Virtualization Virtualization is the process of taking all of the hardware of a physical computer and replicating it in software alone. This means that a single computer can “host” or run several “virtual” computers. These virtual computers can run anywhere - including at a vendor’s location. Some companies refer to this as Cloud Computing since the hardware is operated and maintained elsewhere. IaaS The more detailed definition of this type of computing is called Infrastructure as a Service (Iaas) since it removes the need for you to maintain hardware at your organization. The operating system, drivers, and all the other software required to run an application are still under your control and your responsibility to license, patch, and scale. Microsoft has an offering in this space called Hyper-V, that runs on the Windows operating system. Combined with a hardware hosting vendor and the System Center software to create and deploy Virtual Machines (a process referred to as provisioning), you can create a Cloud environment with full control over all aspects of the machine, including multiple operating systems if you like. Hosting machines and provisioning them at your own buildings is sometimes called a Private Cloud, and hosting them somewhere else is often called a Public Cloud. State-ful and Stateless Programming This paradigm does not create a new, scalable way of computing. It simply moves the hardware away. The reason is that when you limit the Cloud efforts to a Virtual Machine, you are in effect limiting the computing resources to what that single system can provide. This is because much of the software developed in this environment maintains “state” - and that requires a little explanation. “State-ful programming” means that all parts of the computing environment stay connected to each other throughout a compute cycle. The system expects the memory, CPU, storage and network to remain in the same state from the beginning of the process to the end. You can think of this as a telephone conversation - you expect that the other person picks up the phone, listens to you, and talks back all in a single unit of time. In “Stateless” computing the system is designed to allow the different parts of the code to run independently of each other. You can think of this like an e-mail exchange. You compose an e-mail from your system (it has the state when you’re doing that) and then you walk away for a bit to make some coffee. A few minutes later you click the “send” button (the network has the state) and you go to a meeting. The server receives the message and stores it on a mail program’s database (the mail server has the state now) and continues working on other mail. Finally, the other party logs on to their mail client and reads the mail (the other user has the state) and responds to it and so on. These events might be separated by milliseconds or even days, but the system continues to operate. The entire process doesn’t maintain the state, each component does. This is the exact concept behind coding for Windows Azure. The stateless programming model allows amazing rates of scale, since the message (think of the e-mail) can be broken apart by multiple programs and worked on in parallel (like when the e-mail goes to hundreds of users), and only the order of re-assembling the work is important to consider. For the exact same reason, if the system makes copies of those running programs as Windows Azure does, you have built-in redundancy and recovery. It’s just built into the design. The Difference Between Infrastructure Designs and Platform Designs When you simply take a physical server running software and virtualize it either privately or publicly, you haven’t done anything to allow the code to scale or have recovery. That all has to be handled by adding more code and more Virtual Machines that have a slight lag in maintaining the running state of the system. Add more machines and you get more lag, so the scale is limited. This is the primary limitation with IaaS. It’s also not as easy to deploy these VM’s, and more importantly, you’re often charged on a longer basis to remove them. your agility in IaaS is more limited. Windows Azure is a Platform - meaning that you get objects you can code against. The code you write runs on multiple nodes with multiple copies, and it all works because of the magic of Stateless programming. you don’t worry, or even care, about what is running underneath. It could be Windows (and it is in fact a type of Windows Server), Linux, or anything else - but that' isn’t what you want to manage, monitor, maintain or license. You don’t want to deploy an operating system - you want to deploy an application. You want your code to run, and you don’t care how it does that. Another benefit to PaaS is that you can ask for hundreds or thousands of new nodes of computing power - there’s no provisioning, it just happens. And you can stop using them quicker - and the base code for your application does not have to change to make this happen. Windows Azure Roles and Their Use If you need your code to have a user interface, in Visual Studio you add a Web Role to your project, and if the code needs to do work that doesn’t involve a user interface you can add a Worker Role. They are just containers that act a certain way. I’ll provide more detail on those later. Note: That’s a general description, so it’s not entirely accurate, but it’s accurate enough for this discussion. So now we’re back to that VM Role. Because of the name, some have mistakenly thought that you can take a Virtual Machine running, say Linux, and deploy it to Windows Azure using this Role. But you can’t. That’s not what it is designed for at all. If you do need that kind of deployment, you should look into Hyper-V and System Center to create the Private or Public Infrastructure as a Service. What the VM Role is actually designed to do is to allow you to have a great deal of control over the system where your code will run. Let’s take an example. You’ve heard about Windows Azure, and Platform programming. You’re convinced it’s the right way to code. But you have a lot of things you’ve written in another way at your company. Re-writing all of your code to take advantage of Windows Azure will take a long time. Or perhaps you have a certain version of Apache Web Server that you need for your code to work. In both cases, you think you can (or already have) code the the software to be “Stateless”, you just need more control over the place where the code runs. That’s the place where a VM Role makes sense. Recap Virtualizing servers alone has limitations of scale, availability and recovery. Microsoft’s offering in this area is Hyper-V and System Center, not the VM Role. The VM Role is still used for running Stateless code, just like the Web and Worker Roles, with the exception that it allows you more control over the environment of where that code runs.

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  • MySQL – Scalability on Amazon RDS: Scale out to multiple RDS instances

    - by Pinal Dave
    Today, I’d like to discuss getting better MySQL scalability on Amazon RDS. The question of the day: “What can you do when a MySQL database needs to scale write-intensive workloads beyond the capabilities of the largest available machine on Amazon RDS?” Let’s take a look. In a typical EC2/RDS set-up, users connect to app servers from their mobile devices and tablets, computers, browsers, etc.  Then app servers connect to an RDS instance (web/cloud services) and in some cases they might leverage some read-only replicas.   Figure 1. A typical RDS instance is a single-instance database, with read replicas.  This is not very good at handling high write-based throughput. As your application becomes more popular you can expect an increasing number of users, more transactions, and more accumulated data.  User interactions can become more challenging as the application adds more sophisticated capabilities. The result of all this positive activity: your MySQL database will inevitably begin to experience scalability pressures. What can you do? Broadly speaking, there are four options available to improve MySQL scalability on RDS. 1. Larger RDS Instances – If you’re not already using the maximum available RDS instance, you can always scale up – to larger hardware.  Bigger CPUs, more compute power, more memory et cetera. But the largest available RDS instance is still limited.  And they get expensive. “High-Memory Quadruple Extra Large DB Instance”: 68 GB of memory 26 ECUs (8 virtual cores with 3.25 ECUs each) 64-bit platform High I/O Capacity Provisioned IOPS Optimized: 1000Mbps 2. Provisioned IOPs – You can get provisioned IOPs and higher throughput on the I/O level. However, there is a hard limit with a maximum instance size and maximum number of provisioned IOPs you can buy from Amazon and you simply cannot scale beyond these hardware specifications. 3. Leverage Read Replicas – If your application permits, you can leverage read replicas to offload some reads from the master databases. But there are a limited number of replicas you can utilize and Amazon generally requires some modifications to your existing application. And read-replicas don’t help with write-intensive applications. 4. Multiple Database Instances – Amazon offers a fourth option: “You can implement partitioning,thereby spreading your data across multiple database Instances” (Link) However, Amazon does not offer any guidance or facilities to help you with this. “Multiple database instances” is not an RDS feature.  And Amazon doesn’t explain how to implement this idea. In fact, when asked, this is the response on an Amazon forum: Q: Is there any documents that describe the partition DB across multiple RDS? I need to use DB with more 1TB but exist a limitation during the create process, but I read in the any FAQ that you need to partition database, but I don’t find any documents that describe it. A: “DB partitioning/sharding is not an official feature of Amazon RDS or MySQL, but a technique to scale out database by using multiple database instances. The appropriate way to split data depends on the characteristics of the application or data set. Therefore, there is no concrete and specific guidance.” So now what? The answer is to scale out with ScaleBase. Amazon RDS with ScaleBase: What you get – MySQL Scalability! ScaleBase is specifically designed to scale out a single MySQL RDS instance into multiple MySQL instances. Critically, this is accomplished with no changes to your application code.  Your application continues to “see” one database.   ScaleBase does all the work of managing and enforcing an optimized data distribution policy to create multiple MySQL instances. With ScaleBase, data distribution, transactions, concurrency control, and two-phase commit are all 100% transparent and 100% ACID-compliant, so applications, services and tooling continue to interact with your distributed RDS as if it were a single MySQL instance. The result: now you can cost-effectively leverage multiple MySQL RDS instance to scale out write-intensive workloads to an unlimited number of users, transactions, and data. Amazon RDS with ScaleBase: What you keep – Everything! And how does this change your Amazon environment? 1. Keep your application, unchanged – There is no change your application development life-cycle at all.  You still use your existing development tools, frameworks and libraries.  Application quality assurance and testing cycles stay the same. And, critically, you stay with an ACID-compliant MySQL environment. 2. Keep your RDS value-added services – The value-added services that you rely on are all still available. Amazon will continue to handle database maintenance and updates for you. You can still leverage High Availability via Multi A-Z.  And, if it benefits youra application throughput, you can still use read replicas. 3. Keep your RDS administration – Finally the RDS monitoring and provisioning tools you rely on still work as they did before. With your one large MySQL instance, now split into multiple instances, you can actually use less expensive, smallersmaller available RDS hardware and continue to see better database performance. Conclusion Amazon RDS is a tremendous service, but it doesn’t offer solutions to scale beyond a single MySQL instance. Larger RDS instances get more expensive.  And when you max-out on the available hardware, you’re stuck.  Amazon recommends scaling out your single instance into multiple instances for transaction-intensive apps, but offers no services or guidance to help you. This is where ScaleBase comes in to save the day. It gives you a simple and effective way to create multiple MySQL RDS instances, while removing all the complexities typically caused by “DIY” sharding andwith no changes to your applications . With ScaleBase you continue to leverage the AWS/RDS ecosystem: commodity hardware and value added services like read replicas, multi A-Z, maintenance/updates and administration with monitoring tools and provisioning. SCALEBASE ON AMAZON If you’re curious to try ScaleBase on Amazon, it can be found here – Download NOW. Reference: Pinal Dave (http://blog.sqlauthority.com)Filed under: MySQL, PostADay, SQL, SQL Authority, SQL Optimization, SQL Performance, SQL Query, SQL Server, SQL Tips and Tricks, T SQL

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  • Networking in VirtualBox

    - by Fat Bloke
    Networking in VirtualBox is extremely powerful, but can also be a bit daunting, so here's a quick overview of the different ways you can setup networking in VirtualBox, with a few pointers as to which configurations should be used and when. VirtualBox allows you to configure up to 8 virtual NICs (Network Interface Controllers) for each guest vm (although only 4 are exposed in the GUI) and for each of these NICs you can configure: Which virtualized NIC-type is exposed to the Guest. Examples include: Intel PRO/1000 MT Server (82545EM),  AMD PCNet FAST III (Am79C973, the default) or  a Paravirtualized network adapter (virtio-net). How the NIC operates with respect to your Host's physical networking. The main modes are: Network Address Translation (NAT) Bridged networking Internal networking Host-only networking NAT with Port-forwarding The choice of NIC-type comes down to whether the guest has drivers for that NIC.  VirtualBox, suggests a NIC based on the guest OS-type that you specify during creation of the vm, and you rarely need to modify this. But the choice of networking mode depends on how you want to use your vm (client or server) and whether you want other machines on your network to see it. So let's look at each mode in a bit more detail... Network Address Translation (NAT) This is the default mode for new vm's and works great in most situations when the Guest is a "client" type of vm. (i.e. most network connections are outbound). Here's how it works: When the guest OS boots,  it typically uses DHCP to get an IP address. VirtualBox will field this DHCP request and tell the guest OS its assigned IP address and the gateway address for routing outbound connections. In this mode, every vm is assigned the same IP address (10.0.2.15) because each vm thinks they are on their own isolated network. And when they send their traffic via the gateway (10.0.2.2) VirtualBox rewrites the packets to make them appear as though they originated from the Host, rather than the Guest (running inside the Host). This means that the Guest will work even as the Host moves from network to network (e.g. laptop moving between locations), and from wireless to wired connections too. However, how does another computer initiate a connection into a Guest?  e.g. connecting to a web server running in the Guest. This is not (normally) possible using NAT mode as there is no route into the Guest OS. So for vm's running servers we need a different networking mode.... Bridged Networking Bridged Networking is used when you want your vm to be a full network citizen, i.e. to be an equal to your host machine on the network. In this mode, a virtual NIC is "bridged" to a physical NIC on your host, like this: The effect of this is that each VM has access to the physical network in the same way as your host. It can access any service on the network such as external DHCP services, name lookup services, and routing information just as the host does. Logically, the network looks like this: The downside of this mode is that if you run many vm's you can quickly run out of IP addresses or your network administrator gets fed up with you asking for statically assigned IP addresses. Secondly, if your host has multiple physical NICs (e.g. Wireless and Wired) you must reconfigure the bridge when your host jumps networks.  Hmm, so what if you want to run servers in vm's but don't want to involve your network administrator? Maybe one of the next 2 modes is for you... Internal Networking When you configure one or more vm's to sit on an Internal network, VirtualBox ensures that all traffic on that network stays within the host and is only visible to vm's on that virtual network. Configuration looks like this: The internal network ( in this example "intnet" ) is a totally isolated network and so is very "quiet". This is good for testing when you need a separate, clean network, and you can create sophisticated internal networks with vm's that provide their own services to the internal network. (e.g. Active Directory, DHCP, etc). Note that not even the Host is a member of the internal network, but this mode allows vm's to function even when the Host is not connected to a network (e.g. on a plane). Note that in this mode, VirtualBox provides no "convenience" services such as DHCP, so your machines must be statically configured or one of the vm's needs to provide a DHCP/Name service. Multiple internal networks are possible and you can configure vm's to have multiple NICs to sit across internal and other network modes and thereby provide routes if needed. But all this sounds tricky. What if you want an Internal Network that the host participates on with VirtualBox providing IP addresses to the Guests? Ah, then for this, you might want to consider Host-only Networking... Host-only Networking Host-only Networking is like Internal Networking in that you indicate which network the Guest sits on, in this case, "vboxnet0": All vm's sitting on this "vboxnet0" network will see each other, and additionally, the host can see these vm's too. However, other external machines cannot see Guests on this network, hence the name "Host-only". Logically, the network looks like this: This looks very similar to Internal Networking but the host is now on "vboxnet0" and can provide DHCP services. To configure how a Host-only network behaves, look in the VirtualBox Manager...Preferences...Network dialog: Port-Forwarding with NAT Networking Now you may think that we've provided enough modes here to handle every eventuality but here's just one more... What if you cart around a mobile-demo or dev environment on, say, a laptop and you have one or more vm's that you need other machines to connect into? And you are continually hopping onto different (customer?) networks. In this scenario: NAT - won't work because external machines need to connect in. Bridged - possibly an option, but does your customer want you eating IP addresses and can your software cope with changing networks? Internal - we need the vm(s) to be visible on the network, so this is no good. Host-only - same problem as above, we want external machines to connect in to the vm's. Enter Port-forwarding to save the day! Configure your vm's to use NAT networking; Add Port Forwarding rules; External machines connect to "host":"port number" and connections are forwarded by VirtualBox to the guest:port number specified. For example, if your vm runs a web server on port 80, you could set up rules like this:  ...which reads: "any connections on port 8080 on the Host will be forwarded onto this vm's port 80".  This provides a mobile demo system which won't need re-configuring every time you open your laptop lid. Summary VirtualBox has a very powerful set of options allowing you to set up almost any configuration your heart desires. For more information, check out the VirtualBox User Manual on Virtual Networking. -FB 

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  • MIX 2010 Covert Operations Day 2 Silverlight + Windows 7 Phone

    - by GeekAgilistMercenary
    Left the Circus Circus and headed to the geek circus at Mandalay Bay.  Got in, got some breakfast, met a few more people and headed to the keynote. Upon arriving the crew I was hanging with at the event; Erik Mork, Beth Murray, and Brian Henderson and I were entertained with several other thousand geeks by the wicked yo-yoing. The first video demo of something was of Bing Maps and various aspects of Microsoft Research integrated together.  Namely the pictures, put in place, on real 3d element maps of various environments. Silverlight Scott Guthrie, as one would guess, kicked off the keynote.  His first point was that user experience has become a priority at Microsoft.  This can be seen by any observant soul with the release and push of Expression, Silverlight, and the other tools.  This is even more apparent when one takes note of Microsoft bringing in people that can actually do good design and putting them at the forefront. The next thing Scott brought up was a few key points about Silverlight.  Currently Silverlight is a little over 2 years old and has achieved a pretty solid 60% penetration.  Silverlight has all sorts of capabilities that have been developed and are now provided as open source including;  ad injection, smoothing, playback editing, and more.  Another thing he showed, which really struck me as awesome being in the analytics space, was the Olympics and a quick glimpse of the ad statistics, viewer experience, video playback performance, audience trends, and overall viewer participation.  All of it rendered in Silverlight in beautiful detail. The key piece of Scott's various points were all punctuated with the fact that all of this code is available as open source.  Not only is Microsoft really delving into this design element of things, they're getting involved in the right ways. One of the last points I'll bring up about Silverlight 4 is the ability to have HD video on a monitor, and an entirely different activity being done on the other monitor, effectively making Silverlight the only RIA framework that supports multi-monitor support.  Overall, Silverlight is continuing to impress – providing superior capabilities tit-for-tat with the competition. Windows 7 Phone The Windows 7 Phone has 3 primary buttons (yes, more than the iPhone, don't let your mind explode!!).  Start, Search, and Back control all of the needed functionality of the phone.  At the same time, of course, there is the multi-touch, touch, and other interactive abilities of the interface.  The intent, once start is pressed is to have all the information that a phone owner wants displayed immediately.  Avoiding the scrolling through pages of apps or rolling a ball to get through multitudes of other non-interactive phone interfaces.  The Windows 7 Phone simply has the data right in front of you, basically a phone dashboard.  From there it is easy to dive into the interactive areas of the phone. Each area of the interface of the phone is broken into hubs.  These hubs include applications, data, and other things based on a relative basis.  This basis being determined by the user.  These applications interact on many other levels, and form a kind of relationship between each other adding more and more meta-data to the phone user, their interactions between the applications, and of course the social element of their interactions on the phone.  This makes this phone a practical must have for a marketer involved in social media.  The level of wired together interaction is massive, and of course, if you've seen Office Outlook 2010 you know that the power that is pulled into the phone by being tied to Outlook is massive. Joe Belfiore also showed several UI & specifically UX elements of the phone interface that allows paging to be instinctual by simple clipped items, flipping page to page, and other excellent user experience advances for phone devices.  Belfiore's also showed how his people hub had a massive list of people, with pictures, all from various different social networks and other associated relations.  The rendering, speed, and viewing of these people's, their pictures, their social network information, and other characteristics was smooth and in some situations unbelievably rendered.  This demo showed some of the great power of the beta phone, which isn't even as powerful as the planned end device. Joe finished up by jumping into the music, videos, and other media with the Zune Component of the Windows 7 Mobile Phone.  This was all good stuff, but I'll get to what really sold me on the media element in a moment. When Joe was done, Scott Guthrie stepped back up to walk through building a Windows 7 Mobile Phone.  This is were I have to give serious props.  He built this application, in Visual Studio 2010, in front of 2000+ people.  That was cool, but what really was amazing that he build the application in about 2 minutes.  The IDE, side by side design that is standard in Visual Studio is light years ahead of x-Code or any of the iPhone IDEs.  The Windows 7 Mobile System, if it can get market penetration, poses a technologically superior development and phone platform over anything on the market right now.  The biggest problem with the phone, is it just isn't available yet.  I personally can't wait for a chance to build some apps for the new Windows Phone. Netflix, I May Start Up an Account Again! When I get my Windows 7 Phone device, I am absolutely getting a Netflix account again.  The Vertigo crew, as I wrote on Twitter "#MIX10 Props @seesharp on @netflix demo", displayed an application on the phone for Netflix that actually ran HD Video of Rescue Me (with Dennis Leary).  The video played back smooth as it would on a dedicated computer, I was instantly sold.  So this didn't actually sell me on the phone, because I'm already sold, but it did sell me whole heartedly on the media capabilities of the pending phone. Anyway, I try not to do this but I may double post today.  Lunch is over and I'm off to another session very near and dear to the heart of my occupation, Analytics Tracking.  Stay tuned and I should have that post up by the end of the day. Original Post – Check out my other blog for even more technical ramblings and reads.

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  • Can't log in to GNOME after upgrade (raring -> saucy)

    - by x-yuri
    I've just upgraded my ubuntu (raring to saucy) and I now can't log in to GNOME. As opposed to virtual consoles (Ctrl-Alt-F1, for example). I set it up to log in automatically. But it asks for password now. I type in the password, press Enter, the screen blinks and here I am again at the login screen. Then I looked into /var/log/Xorg.0.log: [ 33.956] Initializing built-in extension DRI2 [ 33.956] (II) LoadModule: "glx" [ 33.956] (II) Loading /usr/lib/xorg/modules/extensions/libglx.so [ 33.956] (II) Module glx: vendor="X.Org Foundation" [ 33.956] compiled for 1.14.3, module version = 1.0.0 [ 33.956] ABI class: X.Org Server Extension, version 7.0 [ 33.956] (==) AIGLX enabled [ 33.956] Loading extension GLX [ 33.956] (==) Matched fglrx as autoconfigured driver 0 [ 33.956] (==) Matched ati as autoconfigured driver 1 [ 33.956] (==) Matched fglrx as autoconfigured driver 2 [ 33.956] (==) Matched ati as autoconfigured driver 3 [ 33.956] (==) Matched vesa as autoconfigured driver 4 [ 33.956] (==) Matched modesetting as autoconfigured driver 5 [ 33.956] (==) Matched fbdev as autoconfigured driver 6 [ 33.956] (==) Assigned the driver to the xf86ConfigLayout [ 33.956] (II) LoadModule: "fglrx" [ 33.957] (WW) Warning, couldn't open module fglrx [ 33.957] (II) UnloadModule: "fglrx" [ 33.957] (II) Unloading fglrx [ 33.957] (EE) Failed to load module "fglrx" (module does not exist, 0) [ 33.957] (II) LoadModule: "ati" [ 33.957] (WW) Warning, couldn't open module ati [ 33.957] (II) UnloadModule: "ati" [ 33.957] (II) Unloading ati [ 33.957] (EE) Failed to load module "ati" (module does not exist, 0) [ 33.957] (II) LoadModule: "vesa" [ 33.957] (II) Loading /usr/lib/xorg/modules/drivers/vesa_drv.so [ 33.957] (II) Module vesa: vendor="X.Org Foundation" [ 33.957] compiled for 1.14.1, module version = 2.3.2 [ 33.957] Module class: X.Org Video Driver [ 33.957] ABI class: X.Org Video Driver, version 14.1 [ 33.957] (II) LoadModule: "modesetting" [ 33.957] (II) Loading /usr/lib/xorg/modules/drivers/modesetting_drv.so [ 33.957] (II) Module modesetting: vendor="X.Org Foundation" [ 33.957] compiled for 1.14.1, module version = 0.8.0 [ 33.957] Module class: X.Org Video Driver [ 33.957] ABI class: X.Org Video Driver, version 14.1 [ 33.957] (II) LoadModule: "fbdev" [ 33.957] (II) Loading /usr/lib/xorg/modules/drivers/fbdev_drv.so [ 33.958] (II) Module fbdev: vendor="X.Org Foundation" [ 33.958] compiled for 1.14.1, module version = 0.4.3 [ 33.958] Module class: X.Org Video Driver [ 33.958] ABI class: X.Org Video Driver, version 14.1 [ 33.958] (==) Matched fglrx as autoconfigured driver 0 [ 33.958] (==) Matched ati as autoconfigured driver 1 [ 33.958] (==) Matched fglrx as autoconfigured driver 2 [ 33.958] (==) Matched ati as autoconfigured driver 3 [ 33.958] (==) Matched vesa as autoconfigured driver 4 [ 33.958] (==) Matched modesetting as autoconfigured driver 5 [ 33.958] (==) Matched fbdev as autoconfigured driver 6 [ 33.958] (==) Assigned the driver to the xf86ConfigLayout [ 33.958] (II) LoadModule: "fglrx" [ 33.958] (WW) Warning, couldn't open module fglrx [ 33.958] (II) UnloadModule: "fglrx" [ 33.958] (II) Unloading fglrx [ 33.958] (EE) Failed to load module "fglrx" (module does not exist, 0) [ 33.958] (II) LoadModule: "ati" [ 33.958] (WW) Warning, couldn't open module ati [ 33.958] (II) UnloadModule: "ati" [ 33.958] (II) Unloading ati [ 33.958] (EE) Failed to load module "ati" (module does not exist, 0) [ 33.958] (II) LoadModule: "vesa" [ 33.958] (II) Loading /usr/lib/xorg/modules/drivers/vesa_drv.so [ 33.958] (II) Module vesa: vendor="X.Org Foundation" [ 33.958] compiled for 1.14.1, module version = 2.3.2 [ 33.958] Module class: X.Org Video Driver [ 33.958] ABI class: X.Org Video Driver, version 14.1 [ 33.958] (II) UnloadModule: "vesa" [ 33.958] (II) Unloading vesa [ 33.958] (II) Failed to load module "vesa" (already loaded, 0) [ 33.958] (II) LoadModule: "modesetting" [ 33.959] (II) Loading /usr/lib/xorg/modules/drivers/modesetting_drv.so [ 33.959] (II) Module modesetting: vendor="X.Org Foundation" [ 33.959] compiled for 1.14.1, module version = 0.8.0 [ 33.959] Module class: X.Org Video Driver [ 33.959] ABI class: X.Org Video Driver, version 14.1 [ 33.959] (II) UnloadModule: "modesetting" [ 33.959] (II) Unloading modesetting [ 33.959] (II) Failed to load module "modesetting" (already loaded, 0) [ 33.959] (II) LoadModule: "fbdev" [ 33.959] (II) Loading /usr/lib/xorg/modules/drivers/fbdev_drv.so [ 33.959] (II) Module fbdev: vendor="X.Org Foundation" [ 33.959] compiled for 1.14.1, module version = 0.4.3 [ 33.959] Module class: X.Org Video Driver [ 33.959] ABI class: X.Org Video Driver, version 14.1 [ 33.959] (II) UnloadModule: "fbdev" [ 33.959] (II) Unloading fbdev [ 33.959] (II) Failed to load module "fbdev" (already loaded, 0) [ 33.959] (II) VESA: driver for VESA chipsets: vesa [ 33.959] (II) modesetting: Driver for Modesetting Kernel Drivers: kms [ 33.959] (II) FBDEV: driver for framebuffer: fbdev [ 33.959] (++) using VT number 7 If I install fglrx, it reads: [ 37.152] Initializing built-in extension DRI2 [ 37.152] (II) LoadModule: "glx" [ 37.152] (II) Loading /usr/lib/x86_64-linux-gnu/xorg/extra-modules/modules/extensions/libglx.so [ 37.152] (II) Module glx: vendor="Advanced Micro Devices, Inc." [ 37.152] compiled for 6.9.0, module version = 1.0.0 [ 37.152] Loading extension GLX [ 37.153] (==) Matched fglrx as autoconfigured driver 0 [ 37.153] (==) Matched ati as autoconfigured driver 1 [ 37.153] (==) Matched vesa as autoconfigured driver 2 [ 37.153] (==) Matched modesetting as autoconfigured driver 3 [ 37.153] (==) Matched fbdev as autoconfigured driver 4 [ 37.153] (==) Assigned the driver to the xf86ConfigLayout [ 37.153] (II) LoadModule: "fglrx" [ 37.153] (II) Loading /usr/lib/x86_64-linux-gnu/xorg/extra-modules/modules/drivers/fglrx_drv.so [ 37.168] (II) Module fglrx: vendor="FireGL - AMD Technologies Inc." [ 37.168] compiled for 1.4.99.906, module version = 13.10.10 [ 37.168] Module class: X.Org Video Driver [ 37.168] (II) Loading sub module "fglrxdrm" [ 37.168] (II) LoadModule: "fglrxdrm" [ 37.168] (II) Loading /usr/lib/x86_64-linux-gnu/xorg/extra-modules/modules/linux/libfglrxdrm.so [ 37.169] (II) Module fglrxdrm: vendor="FireGL - AMD Technologies Inc." [ 37.169] compiled for 1.4.99.906, module version = 13.10.10 [ 37.169] (II) LoadModule: "ati" [ 37.169] (WW) Warning, couldn't open module ati [ 37.169] (II) UnloadModule: "ati" [ 37.169] (II) Unloading ati [ 37.169] (EE) Failed to load module "ati" (module does not exist, 0) [ 37.169] (II) LoadModule: "vesa" [ 37.169] (II) Loading /usr/lib/xorg/modules/drivers/vesa_drv.so [ 37.169] (II) Module vesa: vendor="X.Org Foundation" [ 37.169] compiled for 1.14.1, module version = 2.3.2 [ 37.169] Module class: X.Org Video Driver [ 37.169] ABI class: X.Org Video Driver, version 14.1 [ 37.169] (II) LoadModule: "modesetting" [ 37.170] (II) Loading /usr/lib/xorg/modules/drivers/modesetting_drv.so [ 37.170] (II) Module modesetting: vendor="X.Org Foundation" [ 37.170] compiled for 1.14.1, module version = 0.8.0 [ 37.170] Module class: X.Org Video Driver [ 37.170] ABI class: X.Org Video Driver, version 14.1 [ 37.170] (II) LoadModule: "fbdev" [ 37.170] (II) Loading /usr/lib/xorg/modules/drivers/fbdev_drv.so [ 37.170] (II) Module fbdev: vendor="X.Org Foundation" [ 37.170] compiled for 1.14.1, module version = 0.4.3 [ 37.170] Module class: X.Org Video Driver [ 37.170] ABI class: X.Org Video Driver, version 14.1 [ 37.170] (==) Matched fglrx as autoconfigured driver 0 [ 37.170] (==) Matched ati as autoconfigured driver 1 [ 37.170] (==) Matched vesa as autoconfigured driver 2 [ 37.170] (==) Matched modesetting as autoconfigured driver 3 [ 37.170] (==) Matched fbdev as autoconfigured driver 4 [ 37.170] (==) Assigned the driver to the xf86ConfigLayout [ 37.170] (II) LoadModule: "fglrx" [ 37.170] (II) Loading /usr/lib/x86_64-linux-gnu/xorg/extra-modules/modules/drivers/fglrx_drv.so [ 37.170] (II) Module fglrx: vendor="FireGL - AMD Technologies Inc." [ 37.170] compiled for 1.4.99.906, module version = 13.10.10 [ 37.170] Module class: X.Org Video Driver [ 37.170] (II) LoadModule: "ati" [ 37.170] (WW) Warning, couldn't open module ati [ 37.170] (II) UnloadModule: "ati" [ 37.171] (II) Unloading ati [ 37.171] (EE) Failed to load module "ati" (module does not exist, 0) [ 37.171] (II) LoadModule: "vesa" [ 37.171] (II) Loading /usr/lib/xorg/modules/drivers/vesa_drv.so [ 37.171] (II) Module vesa: vendor="X.Org Foundation" [ 37.171] compiled for 1.14.1, module version = 2.3.2 [ 37.171] Module class: X.Org Video Driver [ 37.171] ABI class: X.Org Video Driver, version 14.1 [ 37.171] (II) UnloadModule: "vesa" [ 37.171] (II) Unloading vesa [ 37.171] (II) Failed to load module "vesa" (already loaded, 0) [ 37.171] (II) LoadModule: "modesetting" [ 37.171] (II) Loading /usr/lib/xorg/modules/drivers/modesetting_drv.so [ 37.171] (II) Module modesetting: vendor="X.Org Foundation" [ 37.171] compiled for 1.14.1, module version = 0.8.0 [ 37.171] Module class: X.Org Video Driver [ 37.171] ABI class: X.Org Video Driver, version 14.1 [ 37.171] (II) UnloadModule: "modesetting" [ 37.171] (II) Unloading modesetting [ 37.171] (II) Failed to load module "modesetting" (already loaded, 0) [ 37.171] (II) LoadModule: "fbdev" [ 37.171] (II) Loading /usr/lib/xorg/modules/drivers/fbdev_drv.so [ 37.171] (II) Module fbdev: vendor="X.Org Foundation" [ 37.171] compiled for 1.14.1, module version = 0.4.3 [ 37.171] Module class: X.Org Video Driver [ 37.171] ABI class: X.Org Video Driver, version 14.1 [ 37.171] (II) UnloadModule: "fbdev" [ 37.171] (II) Unloading fbdev [ 37.171] (II) Failed to load module "fbdev" (already loaded, 0) [ 37.171] (II) AMD Proprietary Linux Driver Version Identifier:13.10.10 [ 37.171] (II) AMD Proprietary Linux Driver Release Identifier: UNSUPPORTED-13.101 [ 37.171] (II) AMD Proprietary Linux Driver Build Date: May 23 2013 15:49:35 [ 37.171] (II) VESA: driver for VESA chipsets: vesa [ 37.171] (II) modesetting: Driver for Modesetting Kernel Drivers: kms [ 37.171] (II) FBDEV: driver for framebuffer: fbdev [ 37.171] (++) using VT number 7 I did more installing/removing packages than that. There were a moment when it said: (EE) Failed to load /usr/lib64/xorg/modules/libglamoregl.so: /usr/lib64/xorg/modules/libglamoregl.so: undefined symbol: _glapi_tls_Context Also there is init: not found in ~/.xsession-errors: /usr/sbin/lightdm-session: 5: exec: init: not found Actually, I'm out of ideas. What about you? :)

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  • 6 Ways to Free Up Hard Drive Space Used by Windows System Files

    - by Chris Hoffman
    We’ve previously covered the standard ways to free up space on Windows. But if you have a small solid-state drive and really want more hard space, there are geekier ways to reclaim hard drive space. Not all of these tips are recommended — in fact, if you have more than enough hard drive space, following these tips may actually be a bad idea. There’s a tradeoff to changing all of these settings. Erase Windows Update Uninstall Files Windows allows you to uninstall patches you install from Windows Update. This is helpful if an update ever causes a problem — but how often do you need to uninstall an update, anyway? And will you really ever need to uninstall updates you’ve installed several years ago? These uninstall files are probably just wasting space on your hard drive. A recent update released for Windows 7 allows you to erase Windows Update files from the Windows Disk Cleanup tool. Open Disk Cleanup, click Clean up system files, check the Windows Update Cleanup option, and click OK. If you don’t see this option, run Windows Update and install the available updates. Remove the Recovery Partition Windows computers generally come with recovery partitions that allow you to reset your computer back to its factory default state without juggling discs. The recovery partition allows you to reinstall Windows or use the Refresh and Reset your PC features. These partitions take up a lot of space as they need to contain a complete system image. On Microsoft’s Surface Pro, the recovery partition takes up about 8-10 GB. On other computers, it may be even larger as it needs to contain all the bloatware the manufacturer included. Windows 8 makes it easy to copy the recovery partition to removable media and remove it from your hard drive. If you do this, you’ll need to insert the removable media whenever you want to refresh or reset your PC. On older Windows 7 computers, you could delete the recovery partition using a partition manager — but ensure you have recovery media ready if you ever need to install Windows. If you prefer to install Windows from scratch instead of using your manufacturer’s recovery partition, you can just insert a standard Window disc if you ever want to reinstall Windows. Disable the Hibernation File Windows creates a hidden hibernation file at C:\hiberfil.sys. Whenever you hibernate the computer, Windows saves the contents of your RAM to the hibernation file and shuts down the computer. When it boots up again, it reads the contents of the file into memory and restores your computer to the state it was in. As this file needs to contain much of the contents of your RAM, it’s 75% of the size of your installed RAM. If you have 12 GB of memory, that means this file takes about 9 GB of space. On a laptop, you probably don’t want to disable hibernation. However, if you have a desktop with a small solid-state drive, you may want to disable hibernation to recover the space. When you disable hibernation, Windows will delete the hibernation file. You can’t move this file off the system drive, as it needs to be on C:\ so Windows can read it at boot. Note that this file and the paging file are marked as “protected operating system files” and aren’t visible by default. Shrink the Paging File The Windows paging file, also known as the page file, is a file Windows uses if your computer’s available RAM ever fills up. Windows will then “page out” data to disk, ensuring there’s always available memory for applications — even if there isn’t enough physical RAM. The paging file is located at C:\pagefile.sys by default. You can shrink it or disable it if you’re really crunched for space, but we don’t recommend disabling it as that can cause problems if your computer ever needs some paging space. On our computer with 12 GB of RAM, the paging file takes up 12 GB of hard drive space by default. If you have a lot of RAM, you can certainly decrease the size — we’d probably be fine with 2 GB or even less. However, this depends on the programs you use and how much memory they require. The paging file can also be moved to another drive — for example, you could move it from a small SSD to a slower, larger hard drive. It will be slower if Windows ever needs to use the paging file, but it won’t use important SSD space. Configure System Restore Windows seems to use about 10 GB of hard drive space for “System Protection” by default. This space is used for System Restore snapshots, allowing you to restore previous versions of system files if you ever run into a system problem. If you need to free up space, you could reduce the amount of space allocated to system restore or even disable it entirely. Of course, if you disable it entirely, you’ll be unable to use system restore if you ever need it. You’d have to reinstall Windows, perform a Refresh or Reset, or fix any problems manually. Tweak Your Windows Installer Disc Want to really start stripping down Windows, ripping out components that are installed by default? You can do this with a tool designed for modifying Windows installer discs, such as WinReducer for Windows 8 or RT Se7en Lite for Windows 7. These tools allow you to create a customized installation disc, slipstreaming in updates and configuring default options. You can also use them to remove components from the Windows disc, shrinking the size of the resulting Windows installation. This isn’t recommended as you could cause problems with your Windows installation by removing important features. But it’s certainly an option if you want to make Windows as tiny as possible. Most Windows users can benefit from removing Windows Update uninstallation files, so it’s good to see that Microsoft finally gave Windows 7 users the ability to quickly and easily erase these files. However, if you have more than enough hard drive space, you should probably leave well enough alone and let Windows manage the rest of these settings on its own. Image Credit: Yutaka Tsutano on Flickr     

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  • Scripting Language Sessions at Oracle OpenWorld and MySQL Connect, 2012

    - by cj
    This posts highlights some great scripting language sessions coming up at the Oracle OpenWorld and MySQL Connect conferences. These events are happening in San Francisco from the end of September. You can search for other interesting conference sessions in the Content Catalog. Also check out what is happening at JavaOne in that event's Content Catalog (I haven't included sessions from it in this post.) To find the timeslots and locations of each session, click their respective link and check the "Session Schedule" box on the top right. GEN8431 - General Session: What’s New in Oracle Database Application Development This general session takes a look at what’s been new in the last year in Oracle Database application development tools using the latest generation of database technology. Topics range from Oracle SQL Developer and Oracle Application Express to Java and PHP. (Thomas Kyte - Architect, Oracle) BOF9858 - Meet the Developers of Database Access Services (OCI, ODBC, DRCP, PHP, Python) This session is your opportunity to meet in person the Oracle developers who have built Oracle Database access tools and products such as the Oracle Call Interface (OCI), Oracle C++ Call Interface (OCCI), and Open Database Connectivity (ODBC) drivers; Transparent Application Failover (TAF); Oracle Database Instant Client; Database Resident Connection Pool (DRCP); Oracle Net Services, and so on. The team also works with those who develop the PHP, Ruby, Python, and Perl adapters for Oracle Database. Come discuss with them the features you like, your pains, and new product enhancements in the latest database technology. CON8506 - Syndication and Consolidation: Oracle Database Driver for MySQL Applications This technical session presents a new Oracle Database driver that enables you to run MySQL applications (written in PHP, Perl, C, C++, and so on) against Oracle Database with almost no code change. Use cases for such a driver include application syndication such as interoperability across a relationship database management system, application migration, and database consolidation. In addition, the session covers enhancements in database technology that enable and simplify the migration of third-party databases and applications to and consolidation with Oracle Database. Attend this session to learn more and see a live demo. (Srinath Krishnaswamy - Director, Software Development, Oracle. Kuassi Mensah - Director Product Management, Oracle. Mohammad Lari - Principal Technical Staff, Oracle ) CON9167 - Current State of PHP and MySQL Together, PHP and MySQL power large parts of the Web. The developers of both technologies continue to enhance their software to ensure that developers can be satisfied despite all their changing and growing needs. This session presents an overview of changes in PHP 5.4, which was released earlier this year and shows you various new MySQL-related features available for PHP, from transparent client-side caching to direct support for scaling and high-availability needs. (Johannes Schlüter - SoftwareDeveloper, Oracle) CON8983 - Sharding with PHP and MySQL In deploying MySQL, scale-out techniques can be used to scale out reads, but for scaling out writes, other techniques have to be used. To distribute writes over a cluster, it is necessary to shard the database and store the shards on separate servers. This session provides a brief introduction to traditional MySQL scale-out techniques in preparation for a discussion on the different sharding techniques that can be used with MySQL server and how they can be implemented with PHP. You will learn about static and dynamic sharding schemes, their advantages and drawbacks, techniques for locating and moving shards, and techniques for resharding. (Mats Kindahl - Senior Principal Software Developer, Oracle) CON9268 - Developing Python Applications with MySQL Utilities and MySQL Connector/Python This session discusses MySQL Connector/Python and the MySQL Utilities component of MySQL Workbench and explains how to write MySQL applications in Python. It includes in-depth explanations of the features of MySQL Connector/Python and the MySQL Utilities library, along with example code to illustrate the concepts. Those interested in learning how to expand or build their own utilities and connector features will benefit from the tips and tricks from the experts. This session also provides an opportunity to meet directly with the engineers and provide feedback on your issues and priorities. You can learn what exists today and influence future developments. (Geert Vanderkelen - Software Developer, Oracle) BOF9141 - MySQL Utilities and MySQL Connector/Python: Python Developers, Unite! Come to this lively discussion of the MySQL Utilities component of MySQL Workbench and MySQL Connector/Python. It includes in-depth explanations of the features and dives into the code for those interested in learning how to expand or build their own utilities and connector features. This is an audience-driven session, so put on your best Python shirt and let’s talk about MySQL Utilities and MySQL Connector/Python. (Geert Vanderkelen - Software Developer, Oracle. Charles Bell - Senior Software Developer, Oracle) CON3290 - Integrating Oracle Database with a Social Network Facebook, Flickr, YouTube, Google Maps. There are many social network sites, each with their own APIs for sharing data with them. Most developers do not realize that Oracle Database has base tools for communicating with these sites, enabling all manner of information, including multimedia, to be passed back and forth between the sites. This technical presentation goes through the methods in PL/SQL for connecting to, and then sending and retrieving, all types of data between these sites. (Marcelle Kratochvil - CTO, Piction) CON3291 - Storing and Tuning Unstructured Data and Multimedia in Oracle Database Database administrators need to learn new skills and techniques when the decision is made in their organization to let Oracle Database manage its unstructured data. They will face new scalability challenges. A single row in a table can become larger than a whole database. This presentation covers the techniques a DBA needs for managing the large volume of data in a standard Oracle Database instance. (Marcelle Kratochvil - CTO, Piction) CON3292 - Using PHP, Perl, Visual Basic, Ruby, and Python for Multimedia in Oracle Database These five programming languages are just some of the most popular ones in use at the moment in the marketplace. This presentation details how you can use them to access and retrieve multimedia from Oracle Database. It covers programming techniques and methods for achieving faster development against Oracle Database. (Marcelle Kratochvil - CTO, Piction) UGF5181 - Building Real-World Oracle DBA Tools in Perl Perl is not normally associated with building mission-critical application or DBA tools. Learn why Perl could be a good choice for building your next killer DBA app. This session draws on real-world experience of building DBA tools in Perl, showing the framework and architecture needed to deal with portability, efficiency, and maintainability. Topics include Perl frameworks; Which Comprehensive Perl Archive Network (CPAN) modules are good to use; Perl and CPAN module licensing; Perl and Oracle connectivity; Compiling and deploying your app; An example of what is possible with Perl. (Arjen Visser - CEO & CTO, Dbvisit Software Limited) CON3153 - Perl: A DBA’s and Developer’s Best (Forgotten) Friend This session reintroduces Perl as a language of choice for many solutions for DBAs and developers. Discover what makes Perl so successful and why it is so versatile in our day-to-day lives. Perl can automate all those manual tasks and is truly platform-independent. Perl may not be in the limelight the way other languages are, but it is a remarkable language, it is still very current with ongoing development, and it has amazing online resources. Learn what makes Perl so great (including CPAN), get an introduction to Perl language syntax, find out what you can use Perl for, hear how Oracle uses Perl, discover the best way to learn Perl, and take away a small Perl project challenge. (Arjen Visser - CEO & CTO, Dbvisit Software Limited) CON10332 - Oracle RightNow CX Cloud Service’s Connect PHP API: Intro, What’s New, and Roadmap Connect PHP is a public API that enables developers to build solutions with the Oracle RightNow CX Cloud Service platform. This API is used primarily by developers working within the Oracle RightNow Customer Portal Cloud Service framework who are looking to gain access to data and services hosted by the Oracle RightNow CX Cloud Service platform through a backward-compatible API. Connect for PHP leverages the same data model and services as the Connect Web Services for SOAP API. Come to this session to get an introduction and learn what’s new and what’s coming up. (Mark Rhoads - Senior Principal Applications Engineer, Oracle. Mark Ericson - Sr. Principle Product Manager, Oracle) CON10330 - Oracle RightNow CX Cloud Service APIs and Frameworks Overview Oracle RightNow CX Cloud Service APIs are available in the following areas: desktop UI, Web services, customer portal, PHP, and knowledge. These frameworks provide access to Oracle RightNow CX Cloud Service’s Connect Common Object Model and custom objects. This session provides a broad overview of capabilities in all these areas. (Mark Ericson - Sr. Principle Product Manager, Oracle)

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  • C# Extension Methods - To Extend or Not To Extend...

    - by James Michael Hare
    I've been thinking a lot about extension methods lately, and I must admit I both love them and hate them. They are a lot like sugar, they taste so nice and sweet, but they'll rot your teeth if you eat them too much.   I can't deny that they aren't useful and very handy. One of the major components of the Shared Component library where I work is a set of useful extension methods. But, I also can't deny that they tend to be overused and abused to willy-nilly extend every living type.   So what constitutes a good extension method? Obviously, you can write an extension method for nearly anything whether it is a good idea or not. Many times, in fact, an idea seems like a good extension method but in retrospect really doesn't fit.   So what's the litmus test? To me, an extension method should be like in the movies when a person runs into their twin, separated at birth. You just know you're related. Obviously, that's hard to quantify, so let's try to put a few rules-of-thumb around them.   A good extension method should:     Apply to any possible instance of the type it extends.     Simplify logic and improve readability/maintainability.     Apply to the most specific type or interface applicable.     Be isolated in a namespace so that it does not pollute IntelliSense.     So let's look at a few examples in relation to these rules.   The first rule, to me, is the most important of all. Once again, it bears repeating, a good extension method should apply to all possible instances of the type it extends. It should feel like the long lost relative that should have been included in the original class but somehow was missing from the family tree.    Take this nifty little int extension, I saw this once in a blog and at first I really thought it was pretty cool, but then I started noticing a code smell I couldn't quite put my finger on. So let's look:       public static class IntExtensinos     {         public static int Seconds(int num)         {             return num * 1000;         }           public static int Minutes(int num)         {             return num * 60000;         }     }     This is so you could do things like:       ...     Thread.Sleep(5.Seconds());     ...     proxy.Timeout = 1.Minutes();     ...     Awww, you say, that's cute! Well, that's the problem, it's kitschy and it doesn't always apply (and incidentally you could achieve the same thing with TimeStamp.FromSeconds(5)). It's syntactical candy that looks cool, but tends to rot and pollute the code. It would allow things like:       total += numberOfTodaysOrders.Seconds();     which makes no sense and should never be allowed. The problem is you're applying an extension method to a logical domain, not a type domain. That is, the extension method Seconds() doesn't really apply to ALL ints, it applies to ints that are representative of time that you want to convert to milliseconds.    Do you see what I mean? The two problems, in a nutshell, are that a) Seconds() called off a non-time value makes no sense and b) calling Seconds() off something to pass to something that does not take milliseconds will be off by a factor of 1000 or worse.   Thus, in my mind, you should only ever have an extension method that applies to the whole domain of that type.   For example, this is one of my personal favorites:       public static bool IsBetween<T>(this T value, T low, T high)         where T : IComparable<T>     {         return value.CompareTo(low) >= 0 && value.CompareTo(high) <= 0;     }   This allows you to check if any IComparable<T> is within an upper and lower bound. Think of how many times you type something like:       if (response.Employee.Address.YearsAt >= 2         && response.Employee.Address.YearsAt <= 10)     {     ...     }     Now, you can instead type:       if(response.Employee.Address.YearsAt.IsBetween(2, 10))     {     ...     }     Note that this applies to all IComparable<T> -- that's ints, chars, strings, DateTime, etc -- and does not depend on any logical domain. In addition, it satisfies the second point and actually makes the code more readable and maintainable.   Let's look at the third point. In it we said that an extension method should fit the most specific interface or type possible. Now, I'm not saying if you have something that applies to enumerables, you create an extension for List, Array, Dictionary, etc (though you may have reasons for doing so), but that you should beware of making things TOO general.   For example, let's say we had an extension method like this:       public static T ConvertTo<T>(this object value)     {         return (T)Convert.ChangeType(value, typeof(T));     }         This lets you do more fluent conversions like:       double d = "5.0".ConvertTo<double>();     However, if you dig into Reflector (LOVE that tool) you will see that if the type you are calling on does not implement IConvertible, what you convert to MUST be the exact type or it will throw an InvalidCastException. Now this may or may not be what you want in this situation, and I leave that up to you. Things like this would fail:       object value = new Employee();     ...     // class cast exception because typeof(IEmployee) != typeof(Employee)     IEmployee emp = value.ConvertTo<IEmployee>();       Yes, that's a downfall of working with Convertible in general, but if you wanted your fluent interface to be more type-safe so that ConvertTo were only callable on IConvertibles (and let casting be a manual task), you could easily make it:         public static T ConvertTo<T>(this IConvertible value)     {         return (T)Convert.ChangeType(value, typeof(T));     }         This is what I mean by choosing the best type to extend. Consider that if we used the previous (object) version, every time we typed a dot ('.') on an instance we'd pull up ConvertTo() whether it was applicable or not. By filtering our extension method down to only valid types (those that implement IConvertible) we greatly reduce our IntelliSense pollution and apply a good level of compile-time correctness.   Now my fourth rule is just my general rule-of-thumb. Obviously, you can make extension methods as in-your-face as you want. I included all mine in my work libraries in its own sub-namespace, something akin to:       namespace Shared.Core.Extensions { ... }     This is in a library called Shared.Core, so just referencing the Core library doesn't pollute your IntelliSense, you have to actually do a using on Shared.Core.Extensions to bring the methods in. This is very similar to the way Microsoft puts its extension methods in System.Linq. This way, if you want 'em, you use the appropriate namespace. If you don't want 'em, they won't pollute your namespace.   To really make this work, however, that namespace should only include extension methods and subordinate types those extensions themselves may use. If you plant other useful classes in those namespaces, once a user includes it, they get all the extensions too.   Also, just as a personal preference, extension methods that aren't simply syntactical shortcuts, I like to put in a static utility class and then have extension methods for syntactical candy. For instance, I think it imaginable that any object could be converted to XML:       namespace Shared.Core     {         // A collection of XML Utility classes         public static class XmlUtility         {             ...             // Serialize an object into an xml string             public static string ToXml(object input)             {                 var xs = new XmlSerializer(input.GetType());                   // use new UTF8Encoding here, not Encoding.UTF8. The later includes                 // the BOM which screws up subsequent reads, the former does not.                 using (var memoryStream = new MemoryStream())                 using (var xmlTextWriter = new XmlTextWriter(memoryStream, new UTF8Encoding()))                 {                     xs.Serialize(xmlTextWriter, input);                     return Encoding.UTF8.GetString(memoryStream.ToArray());                 }             }             ...         }     }   I also wanted to be able to call this from an object like:       value.ToXml();     But here's the problem, if i made this an extension method from the start with that one little keyword "this", it would pop into IntelliSense for all objects which could be very polluting. Instead, I put the logic into a utility class so that users have the choice of whether or not they want to use it as just a class and not pollute IntelliSense, then in my extensions namespace, I add the syntactical candy:       namespace Shared.Core.Extensions     {         public static class XmlExtensions         {             public static string ToXml(this object value)             {                 return XmlUtility.ToXml(value);             }         }     }   So now it's the best of both worlds. On one hand, they can use the utility class if they don't want to pollute IntelliSense, and on the other hand they can include the Extensions namespace and use as an extension if they want. The neat thing is it also adheres to the Single Responsibility Principle. The XmlUtility is responsible for converting objects to XML, and the XmlExtensions is responsible for extending object's interface for ToXml().

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  • Use BGInfo to Build a Database of System Information of Your Network Computers

    - by Sysadmin Geek
    One of the more popular tools of the Sysinternals suite among system administrators is BGInfo which tacks real-time system information to your desktop wallpaper when you first login. For obvious reasons, having information such as system memory, available hard drive space and system up time (among others) right in front of you is very convenient when you are managing several systems. A little known feature about this handy utility is the ability to have system information automatically saved to a SQL database or some other data file. With a few minutes of setup work you can easily configure BGInfo to record system information of all your network computers in a centralized storage location. You can then use this data to monitor or report on these systems however you see fit. BGInfo Setup If you are familiar with BGInfo, you can skip this section. However, if you have never used this tool, it takes just a few minutes to setup in order to capture the data you are looking for. When you first open BGInfo, a timer will be counting down in the upper right corner. Click the countdown button to keep the interface up so we can edit the settings. Now edit the information you want to capture from the available fields on the right. Since all the output will be redirected to a central location, don’t worry about configuring the layout or formatting. Configuring the Storage Database BGInfo supports the ability to store information in several database formats: SQL Server Database, Access Database, Excel and Text File. To configure this option, open File > Database. Using a Text File The simplest, and perhaps most practical, option is to store the BGInfo data in a comma separated text file. This format allows for the file to be opened in Excel or imported into a database. To use a text file or any other file system type (Excel or MS Access), simply provide the UNC to the respective file. The account running the task to write to this file will need read/write access to both the share and NTFS file permissions. When using a text file, the only option is to have BGInfo create a new entry each time the capture process is run which will add a new line to the respective CSV text file. Using a SQL Database If you prefer to have the data dropped straight into a SQL Server database, BGInfo support this as well. This requires a bit of additional configuration, but overall it is very easy. The first step is to create a database where the information will be stored. Additionally, you will want to create a user account to fill data into this table (and this table only). For your convenience, this script creates a new database and user account (run this as Administrator on your SQL Server machine): @SET Server=%ComputerName%.@SET Database=BGInfo@SET UserName=BGInfo@SET Password=passwordSQLCMD -S “%Server%” -E -Q “Create Database [%Database%]“SQLCMD -S “%Server%” -E -Q “Create Login [%UserName%] With Password=N’%Password%’, DEFAULT_DATABASE=[%Database%], CHECK_EXPIRATION=OFF, CHECK_POLICY=OFF”SQLCMD -S “%Server%” -E -d “%Database%” -Q “Create User [%UserName%] For Login [%UserName%]“SQLCMD -S “%Server%” -E -d “%Database%” -Q “EXEC sp_addrolemember N’db_owner’, N’%UserName%’” Note the SQL user account must have ‘db_owner’ permissions on the database in order for BGInfo to work correctly. This is why you should have a SQL user account specifically for this database. Next, configure BGInfo to connect to this database by clicking on the SQL button. Fill out the connection properties according to your database settings. Select the option of whether or not to only have one entry per computer or keep a history of each system. The data will then be dropped directly into a table named “BGInfoTable” in the respective database.   Configure User Desktop Options While the primary function of BGInfo is to alter the user’s desktop by adding system info as part of the wallpaper, for our use here we want to leave the user’s wallpaper alone so this process runs without altering any of the user’s settings. Click the Desktops button. Configure the Wallpaper modifications to not alter anything.   Preparing the Deployment Now we are all set for deploying the configuration to the individual machines so we can start capturing the system data. If you have not done so already, click the Apply button to create the first entry in your data repository. If all is configured correctly, you should be able to open your data file or database and see the entry for the respective machine. Now click the File > Save As menu option and save the configuration as “BGInfoCapture.bgi”.   Deploying to Client Machines Deployment to the respective client machines is pretty straightforward. No installation is required as you just need to copy the BGInfo.exe and the BGInfoCapture.bgi to each machine and place them in the same directory. Once in place, just run the command: BGInfo.exe BGInfoCapture.bgi /Timer:0 /Silent /NoLicPrompt Of course, you probably want to schedule the capture process to run on a schedule. This command creates a Scheduled Task to run the capture process at 8 AM every morning and assumes you copied the required files to the root of your C drive: SCHTASKS /Create /SC DAILY /ST 08:00 /TN “System Info” /TR “C:\BGInfo.exe C:\BGInfoCapture.bgi /Timer:0 /Silent /NoLicPrompt” Adjust as needed, but the end result is the scheduled task command should look something like this:   Download BGInfo from Sysinternals Latest Features How-To Geek ETC How To Create Your Own Custom ASCII Art from Any Image How To Process Camera Raw Without Paying for Adobe Photoshop How Do You Block Annoying Text Message (SMS) Spam? How to Use and Master the Notoriously Difficult Pen Tool in Photoshop HTG Explains: What Are the Differences Between All Those Audio Formats? 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  • SQL SERVER – Core Concepts – Elasticity, Scalability and ACID Properties – Exploring NuoDB an Elastically Scalable Database System

    - by pinaldave
    I have been recently exploring Elasticity and Scalability attributes of databases. You can see that in my earlier blog posts about NuoDB where I wanted to look at Elasticity and Scalability concepts. The concepts are very interesting, and intriguing as well. I have discussed these concepts with my friend Joyti M and together we have come up with this interesting read. The goal of this article is to answer following simple questions What is Elasticity? What is Scalability? How ACID properties vary from NOSQL Concepts? What are the prevailing problems in the current database system architectures? Why is NuoDB  an innovative and welcome change in database paradigm? Elasticity This word’s original form is used in many different ways and honestly it does do a decent job in holding things together over the years as a person grows and contracts. Within the tech world, and specifically related to software systems (database, application servers), it has come to mean a few things - allow stretching of resources without reaching the breaking point (on demand). What are resources in this context? Resources are the usual suspects – RAM/CPU/IO/Bandwidth in the form of a container (a process or bunch of processes combined as modules). When it is about increasing resources the simplest idea which comes to mind is the addition of another container. Another container means adding a brand new physical node. When it is about adding a new node there are two questions which comes to mind. 1) Can we add another node to our software system? 2) If yes, does adding new node cause downtime for the system? Let us assume we have added new node, let us see what the new needs of the system are when a new node is added. Balancing incoming requests to multiple nodes Synchronization of a shared state across multiple nodes Identification of “downstate” and resolution action to bring it to “upstate” Well, adding a new node has its advantages as well. Here are few of the positive points Throughput can increase nearly horizontally across the node throughout the system Response times of application will increase as in-between layer interactions will be improved Now, Let us put the above concepts in the perspective of a Database. When we mention the term “running out of resources” or “application is bound to resources” the resources can be CPU, Memory or Bandwidth. The regular approach to “gain scalability” in the database is to look around for bottlenecks and increase the bottlenecked resource. When we have memory as a bottleneck we look at the data buffers, locks, query plans or indexes. After a point even this is not enough as there needs to be an efficient way of managing such large workload on a “single machine” across memory and CPU bound (right kind of scheduling)  workload. We next move on to either read/write separation of the workload or functionality-based sharing so that we still have control of the individual. But this requires lots of planning and change in client systems in terms of knowing where to go/update/read and for reporting applications to “aggregate the data” in an intelligent way. What we ideally need is an intelligent layer which allows us to do these things without us getting into managing, monitoring and distributing the workload. Scalability In the context of database/applications, scalability means three main things Ability to handle normal loads without pressure E.g. X users at the Y utilization of resources (CPU, Memory, Bandwidth) on the Z kind of hardware (4 processor, 32 GB machine with 15000 RPM SATA drives and 1 GHz Network switch) with T throughput Ability to scale up to expected peak load which is greater than normal load with acceptable response times Ability to provide acceptable response times across the system E.g. Response time in S milliseconds (or agreed upon unit of measure) – 90% of the time The Issue – Need of Scale In normal cases one can plan for the load testing to test out normal, peak, and stress scenarios to ensure specific hardware meets the needs. With help from Hardware and Software partners and best practices, bottlenecks can be identified and requisite resources added to the system. Unfortunately this vertical scale is expensive and difficult to achieve and most of the operational people need the ability to scale horizontally. This helps in getting better throughput as there are physical limits in terms of adding resources (Memory, CPU, Bandwidth and Storage) indefinitely. Today we have different options to achieve scalability: Read & Write Separation The idea here is to do actual writes to one store and configure slaves receiving the latest data with acceptable delays. Slaves can be used for balancing out reads. We can also explore functional separation or sharing as well. We can separate data operations by a specific identifier (e.g. region, year, month) and consolidate it for reporting purposes. For functional separation the major disadvantage is when schema changes or workload pattern changes. As the requirement grows one still needs to deal with scale need in manual ways by providing an abstraction in the middle tier code. Using NOSQL solutions The idea is to flatten out the structures in general to keep all values which are retrieved together at the same store and provide flexible schema. The issue with the stores is that they are compromising on mostly consistency (no ACID guarantees) and one has to use NON-SQL dialect to work with the store. The other major issue is about education with NOSQL solutions. Would one really want to make these compromises on the ability to connect and retrieve in simple SQL manner and learn other skill sets? Or for that matter give up on ACID guarantee and start dealing with consistency issues? Hybrid Deployment – Mac, Linux, Cloud, and Windows One of the challenges today that we see across On-premise vs Cloud infrastructure is a difference in abilities. Take for example SQL Azure – it is wonderful in its concepts of throttling (as it is shared deployment) of resources and ability to scale using federation. However, the same abilities are not available on premise. This is not a mistake, mind you – but a compromise of the sweet spot of workloads, customer requirements and operational SLAs which can be supported by the team. In today’s world it is imperative that databases are available across operating systems – which are a commodity and used by developers of all hues. An Ideal Database Ability List A system which allows a linear scale of the system (increase in throughput with reasonable response time) with the addition of resources A system which does not compromise on the ACID guarantees and require developers to learn new paradigms A system which does not force fit a new way interacting with database by learning Non-SQL dialect A system which does not force fit its mechanisms for providing availability across its various modules. Well NuoDB is the first database which has all of the above abilities and much more. In future articles I will cover my hands-on experience with it. Reference: Pinal Dave (http://blog.SQLAuthority.com) Filed under: PostADay, SQL, SQL Authority, SQL Query, SQL Server, SQL Tips and Tricks, T SQL, Technology Tagged: NuoDB

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  • New ways for backup, recovery and restore of Essbase Block Storage databases – part 2 by Bernhard Kinkel

    - by Alexandra Georgescu
    After discussing in the first part of this article new options in Essbase for the general backup and restore, this second part will deal with the also rather new feature of Transaction Logging and Replay, which was released in version 11.1, enhancing existing restore options. Tip: Transaction logging and replay cannot be used for aggregate storage databases. Please refer to the Oracle Hyperion Enterprise Performance Management System Backup and Recovery Guide (rel. 11.1.2.1). Even if backups are done on a regular, frequent base, subsequent data entries, loads or calculations would not be reflected in a restored database. Activating Transaction Logging could fill that gap and provides you with an option to capture these post-backup transactions for later replay. The following table shows, which are the transactions that could be logged when Transaction Logging is enabled: In order to activate its usage, corresponding statements could be added to the Essbase.cfg file, using the TRANSACTIONLOGLOCATION command. The complete syntax reads: TRANSACTIONLOGLOCATION [ appname [ dbname]] LOGLOCATION NATIVE ?ENABLE | DISABLE Where appname and dbname are optional parameters giving you the chance in combination with the ENABLE or DISABLE command to set Transaction Logging for certain applications or databases or to exclude them from being logged. If only an appname is specified, the setting applies to all databases in that particular application. If appname and dbname are not defined, all applications and databases would be covered. LOGLOCATION specifies the directory to which the log is written, e.g. D:\temp\trlogs. This directory must already exist or needs to be created before using it for log information being written to it. NATIVE is a reserved keyword that shouldn’t be changed. The following example shows how to first enable logging on a more general level for all databases in the application Sample, followed by a disabling statement on a more granular level for only the Basic database in application Sample, hence excluding it from being logged. TRANSACTIONLOGLOCATION Sample Hyperion/trlog/Sample NATIVE ENABLE TRANSACTIONLOGLOCATION Sample Basic Hyperion/trlog/Sample NATIVE DISABLE Tip: After applying changes to the configuration file you must restart the Essbase server in order to initialize the settings. A maybe required replay of logged transactions after restoring a database can be done only by administrators. The following options are available: In Administration Services selecting Replay Transactions on the right-click menu on the database: Here you can select to replay transactions logged after the last replay request was originally executed or after the time of the last restored backup (whichever occurred later) or transactions logged after a specified time. Or you can replay transactions selectively based on a range of sequence IDs, which can be accessed using Display Transactions on the right-click menu on the database: These sequence ID s (0, 1, 2 … 7 in the screenshot below) are assigned to each logged transaction, indicating the order in which the transaction was performed. This helps to ensure the integrity of the restored data after a replay, as the replay of transactions is enforced in the same order in which they were originally performed. So for example a calculation originally run after a data load cannot be replayed before having replayed the data load first. After a transaction is replayed, you can replay only transactions with a greater sequence ID. For example, replaying the transaction with sequence ID of 4 includes all preceding transactions, while afterwards you can only replay transactions with a sequence ID of 5 or greater. Tip: After restoring a database from a backup you should always completely replay all logged transactions, which were executed after the backup, before executing new transactions. But not only the transaction information itself needs to be logged and stored in a specified directory as described above. During transaction logging, Essbase also creates archive copies of data load and rules files in the following default directory: ARBORPATH/app/appname/dbname/Replay These files are then used during the replay of a logged transaction. By default Essbase archives only data load and rules files for client data loads, but in order to specify the type of data to archive when logging transactions you can use the command TRANSACTIONLOGDATALOADARCHIVE as an additional entry in the Essbase.cfg file. The syntax for the statement is: TRANSACTIONLOGDATALOADARCHIVE [appname [dbname]] [OPTION] While to the [appname [dbname]] argument the same applies like before for TRANSACTIONLOGLOCATION, the valid values for the OPTION argument are the following: Make the respective setting for which files copies should be logged, considering from which location transactions are usually taking place. Selecting the NONE option prevents Essbase from saving the respective files and the data load cannot be replayed. In this case you must first manually load the data before you can replay the transactions. Tip: If you use server or SQL data and the data and rules files are not archived in the Replay directory (for example, you did not use the SERVER or SERVER_CLIENT option), Essbase replays the data that is actually in the data source at the moment of the replay, which may or may not be the data that was originally loaded. You can find more detailed information in the following documents: Oracle Hyperion Enterprise Performance Management System Backup and Recovery Guide (rel. 11.1.2.1) Oracle Essbase Online Documentation (rel. 11.1.2.1)) Enterprise Performance Management System Documentation (including previous releases) Or on the Oracle Technology Network. If you are also interested in other new features and smart enhancements in Essbase or Hyperion Planning stay tuned for coming articles or check our training courses and web presentations. You can find general information about offerings for the Essbase and Planning curriculum or other Oracle-Hyperion products here; (please make sure to select your country/region at the top of this page) or in the OU Learning paths section, where Planning, Essbase and other Hyperion products can be found under the Fusion Middleware heading (again, please select the right country/region). Or drop me a note directly: [email protected]. About the Author: Bernhard Kinkel started working for Hyperion Solutions as a Presales Consultant and Consultant in 1998 and moved to Hyperion Education Services in 1999. He joined Oracle University in 2007 where he is a Principal Education Consultant. Based on these many years of working with Hyperion products he has detailed product knowledge across several versions. He delivers both classroom and live virtual courses. His areas of expertise are Oracle/Hyperion Essbase, Oracle Hyperion Planning and Hyperion Web Analysis. Disclaimer: All methods and features mentioned in this article must be considered and tested carefully related to your environment, processes and requirements. As guidance please always refer to the available software documentation. This article does not recommend or advise any explicit action or change, hence the author cannot be held responsible for any consequences due to the use or implementation of these features.

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  • How to Use the Signature Editor in Outlook 2013

    - by Lori Kaufman
    The Signature Editor in Outlook 2013 allows you to create a custom signature from text, graphics, or business cards. We will show you how to use the various features of the Signature Editor to customize your signatures. To open the Signature Editor, click the File tab and select Options on the left side of the Account Information screen. Then, click Mail on the left side of the Options dialog box and click the Signatures button. For more details, refer to one of the articles mentioned above. Changing the font for your signature is pretty self-explanatory. Select the text for which you want to change the font and select the desired font from the drop-down list. You can also set the justification (left, center, right) for each line of text separately. The drop-down list that reads Automatic by default allows you to change the color of the selected text. Click OK to accept your changes and close the Signatures and Stationery dialog box. To see your signature in an email, click Mail on the Navigation Bar. Click New Email on the Home tab. The Message window displays and your default signature is inserted into the body of the email. NOTE: You shouldn’t use fonts that are not common in your signatures. In order for the recipient to see your signature as you intended, the font you choose also needs to be installed on the recipient’s computer. If the font is not installed, the recipient would see a different font, the wrong characters, or even placeholder characters, which are empty square boxes. Close the Message window using the File tab or the X button in the upper, right corner of the Message window. You can save it as a draft if you want, but it’s not necessary. If you decide to use a font that is not common, a better way to do so would be to create a signature as an image, or logo. Create your image or logo in an image editing program making it the exact size you want to use in your signature. Save the image in a file size as small as possible. The .jpg format works well for pictures, the .png format works well for detailed graphics, and the .gif format works well for simple graphics. The .gif format generally produces the smallest files. To insert an image in your signature, open the Signatures and Stationery dialog box again. Either delete the text currently in the editor, if any, or create a new signature. Then, click the image button on the editor’s toolbar. On the Insert Picture dialog box, navigate to the location of your image, select the file, and click Insert. If you want to insert an image from the web, you must enter the full URL for the image in the File name edit box (instead of the local image filename). For example, http://www.somedomain.com/images/signaturepic.gif. If you want to link to the image at the specified URL, you must also select Link to File from the Insert drop-down list to maintain the URL reference. The image is inserted into the Edit signature box. Click OK to accept your changes and close the Signatures and Stationery dialog box. Create a new email message again. You’ll notice the image you inserted into the signature displays in the body of the message. Close the Message window using the File tab or the X button in the upper, right corner of the Message window. You may want to put a link to a webpage or an email link in your signature. To do this, open the Signatures and Stationery dialog box again. Enter the text to display for the link, highlight the text, and click the Hyperlink button on the editor’s toolbar. On the Insert Hyperlink dialog box, select the type of link from the list on the left and enter the webpage, email, or other type of address in the Address edit box. You can change the text that will display in the signature for the link in the Text to display edit box. Click OK to accept your changes and close the dialog box. The link displays in the editor with the default blue, underlined text. Click OK to accept your changes and close the Signatures and Stationery dialog box. Here’s an example of an email message with a link in the signature. Close the Message window using the File tab or the X button in the upper, right corner of the Message window. You can also insert your contact information into your signature as a Business Card. To do so, click Business Card on the editor’s toolbar. On the Insert Business Card dialog box, select the contact you want to insert as a Business Card. Select a size for the Business Card image from the Size drop-down list. Click OK. The Business Card image displays in the Signature Editor. Click OK to accept your changes and close the Signatures and Stationery dialog box. When you insert a Business Card into your signature, the Business Card image displays in the body of the email message and a .vcf file containing your contact information is attached to the email. This .vcf file can be imported into programs like Outlook that support this format. Close the Message window using the File tab or the X button in the upper, right corner of the Message window. You can also insert your Business Card into your signature without the image or without the .vcf file attached. If you want to provide recipients your contact info in a .vcf file, but don’t want to attach it to every email, you can upload the .vcf file to a location on the internet and add a link to the file, such as “Get my vCard,” in your signature. NOTE: If you want to edit your business card, such as applying a different template to it, you must select a different View other than People for your Contacts folder so you can open the full contact editing window.     

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  • Elegance, thy Name is jQuery

    - by SGWellens
    So, I'm browsing though some questions over on the Stack Overflow website and I found a good jQuery question just a few minutes old. Here is a link to it. It was a tough question; I knew that by answering it, I could learn new stuff and reinforce what I already knew: Reading is good, doing is better. Maybe I could help someone in the process too. I cut and pasted the HTML from the question into my Visual Studio IDE and went back to Stack Overflow to reread the question. Dang, someone had already answered it! And it was a great answer. I never even had a chance to start analyzing the issue. Now I know what a one-legged man feels like in an ass-kicking contest. Nevertheless, since the question and answer were so interesting, I decided to dissect them and learn as much as possible. The HTML consisted of some divs separated by h3 headings.  Note the elements are laid out sequentially with no programmatic grouping: <h3 class="heading">Heading 1</h3> <div>Content</div> <div>More content</div> <div>Even more content</div><h3 class="heading">Heading 2</h3> <div>some content</div> <div>some more content</div><h3 class="heading">Heading 3</h3> <div>other content</div></form></body>  The requirement was to wrap a div around each h3 heading and the subsequent divs grouping them into sections. Why? I don't know, I suppose if you screen-scrapped some HTML from another site, you might want to reformat it before displaying it on your own. Anyways… Here is the marvelously, succinct posted answer: $('.heading').each(function(){ $(this).nextUntil('.heading').andSelf().wrapAll('<div class="section">');}); I was familiar with all the parts except for nextUntil and andSelf. But, I'll analyze the whole answer for completeness. I'll do this by rewriting the posted answer in a different style and adding a boat-load of comments: function Test(){ // $Sections is a jQuery object and it will contain three elements var $Sections = $('.heading'); // use each to iterate over each of the three elements $Sections.each(function () { // $this is a jquery object containing the current element // being iterated var $this = $(this); // nextUntil gets the following sibling elements until it reaches // an element with the CSS class 'heading' // andSelf adds in the source element (this) to the collection $this = $this.nextUntil('.heading').andSelf(); // wrap the elements with a div $this.wrapAll('<div class="section" >'); });}  The code here doesn't look nearly as concise and elegant as the original answer. However, unless you and your staff are jQuery masters, during development it really helps to work through algorithms step by step. You can step through this code in the debugger and examine the jQuery objects to make sure one step is working before proceeding on to the next. It's much easier to debug and troubleshoot when each logical coding step is a separate line of code. Note: You may think the original code runs much faster than this version. However, the time difference is trivial: Not enough to worry about: Less than 1 millisecond (tested in IE and FF). Note: You may want to jam everything into one line because it results in less traffic being sent to the client. That is true. However, most Internet servers now compress HTML and JavaScript by stripping out comments and white space (go to Bing or Google and view the source). This feature should be enabled on your server: Let the server compress your code, you don't need to do it. Free Career Advice: Creating maintainable code is Job One—Maximum Priority—The Prime Directive. If you find yourself suddenly transferred to customer support, it may be that the code you are writing is not as readable as it could be and not as readable as it should be. Moving on… I created a CSS class to enhance the results: .section{ background-color: yellow; border: 2px solid black; margin: 5px;} Here is the rendered output before:   …and after the jQuery code runs.   Pretty Cool! But, while playing with this code, the logic of nextUntil began to bother me: What happens in the last section? What stops elements from being collected since there are no more elements with the .heading class? The answer is nothing.  In this case it stopped collecting elements because it was at the end of the page.  But what if there were additional HTML elements? I added an anchor tag and another div to the HTML: <h3 class="heading">Heading 1</h3> <div>Content</div> <div>More content</div> <div>Even more content</div><h3 class="heading">Heading 2</h3> <div>some content</div> <div>some more content</div><h3 class="heading">Heading 3</h3> <div>other content</div><a>this is a link</a><div>unrelated div</div> </form></body> The code as-is will include both the anchor and the unrelated div. This isn't what we want.   My first attempt to correct this used the filter parameter of the nextUntil function: nextUntil('.heading', 'div')  This will only collect div elements. But it merely skipped the anchor tag and it still collected the unrelated div:   The problem is we need a way to tell the nextUntil function when to stop. CSS selectors to the rescue! nextUntil('.heading, a')  This tells nextUntil to stop collecting elements when it gets to an element with a .heading class OR when it gets to an anchor tag. In this case it solved the problem. FYI: The comma operator in a CSS selector allows multiple criteria.   Bingo! One final note, we could have broken the code down even more: We could have replaced the andSelf function here: $this = $this.nextUntil('.heading, a').andSelf(); With this: // get all the following siblings and then add the current item$this = $this.nextUntil('.heading, a');$this.add(this);  But in this case, the andSelf function reads real nice. In my opinion. Here's a link to a jsFiddle if you want to play with it. I hope someone finds this useful Steve Wellens CodeProject

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  • how to use serial port in UDK using windows DLL and DLLBind directive?

    - by Shayan Abbas
    I want to use serial port in UDK, For that purpose i use a windows DLL and DLLBind directive. I have a thread in windows DLL for serial port data recieve event. My problem is: this thread doesn't work properly. Please Help me. below is my code SerialPortDLL Code: // SerialPortDLL.cpp : Defines the exported functions for the DLL application. // #include "stdafx.h" #include "Cport.h" extern "C" { // This is an example of an exported variable //SERIALPORTDLL_API int nSerialPortDLL=0; // This is an example of an exported function. //SERIALPORTDLL_API int fnSerialPortDLL(void) //{ // return 42; //} CPort *sp; __declspec(dllexport) void Open(wchar_t* portName) { sp = new CPort(portName); //MessageBox(0,L"ha ha!!!",L"ha ha",0); //MessageBox(0,portName,L"ha ha",0); } __declspec(dllexport) void Close() { sp->Close(); MessageBox(0,L"ha ha!!!",L"ha ha",0); } __declspec(dllexport) wchar_t *GetData() { return sp->GetData(); } __declspec(dllexport) unsigned int GetDSR() { return sp->getDSR(); } __declspec(dllexport) unsigned int GetCTS() { return sp->getCTS(); } __declspec(dllexport) unsigned int GetRing() { return sp->getRing(); } } CPort class code: #include "stdafx.h" #include "CPort.h" #include "Serial.h" CSerial serial; HANDLE HandleOfThread; LONG lLastError = ERROR_SUCCESS; bool fContinue = true; HANDLE hevtOverlapped; HANDLE hevtStop; OVERLAPPED ov = {0}; //char szBuffer[101] = ""; wchar_t *szBuffer = L""; wchar_t *data = L""; DWORD WINAPI ThreadHandler( LPVOID lpParam ) { // Keep reading data, until an EOF (CTRL-Z) has been received do { MessageBox(0,L"ga ga!!!",L"ga ga",0); //Sleep(10); // Wait for an event lLastError = serial.WaitEvent(&ov); if (lLastError != ERROR_SUCCESS) { //LOG( " Unable to wait for a COM-port event" ); } // Setup array of handles in which we are interested HANDLE ahWait[2]; ahWait[0] = hevtOverlapped; ahWait[1] = hevtStop; // Wait until something happens switch (::WaitForMultipleObjects(sizeof(ahWait)/sizeof(*ahWait),ahWait,FALSE,INFINITE)) { case WAIT_OBJECT_0: { // Save event const CSerial::EEvent eEvent = serial.GetEventType(); // Handle break event if (eEvent & CSerial::EEventBreak) { //LOG( " ### BREAK received ###" ); } // Handle CTS event if (eEvent & CSerial::EEventCTS) { //LOG( " ### Clear to send %s ###", serial.GetCTS() ? "on":"off" ); } // Handle DSR event if (eEvent & CSerial::EEventDSR) { //LOG( " ### Data set ready %s ###", serial.GetDSR() ? "on":"off" ); } // Handle error event if (eEvent & CSerial::EEventError) { switch (serial.GetError()) { case CSerial::EErrorBreak: /*LOG( " Break condition" );*/ break; case CSerial::EErrorFrame: /*LOG( " Framing error" );*/ break; case CSerial::EErrorIOE: /*LOG( " IO device error" );*/ break; case CSerial::EErrorMode: /*LOG( " Unsupported mode" );*/ break; case CSerial::EErrorOverrun: /*LOG( " Buffer overrun" );*/ break; case CSerial::EErrorRxOver: /*LOG( " Input buffer overflow" );*/ break; case CSerial::EErrorParity: /*LOG( " Input parity error" );*/ break; case CSerial::EErrorTxFull: /*LOG( " Output buffer full" );*/ break; default: /*LOG( " Unknown" );*/ break; } } // Handle ring event if (eEvent & CSerial::EEventRing) { //LOG( " ### RING ###" ); } // Handle RLSD/CD event if (eEvent & CSerial::EEventRLSD) { //LOG( " ### RLSD/CD %s ###", serial.GetRLSD() ? "on" : "off" ); } // Handle data receive event if (eEvent & CSerial::EEventRecv) { // Read data, until there is nothing left DWORD dwBytesRead = 0; do { // Read data from the COM-port lLastError = serial.Read(szBuffer,33,&dwBytesRead); if (lLastError != ERROR_SUCCESS) { //LOG( "Unable to read from COM-port" ); } if( dwBytesRead == 33 && szBuffer[0]=='$' ) { // Finalize the data, so it is a valid string szBuffer[dwBytesRead] = '\0'; ////LOG( "\n%s\n", szBuffer ); data = szBuffer; } } while (dwBytesRead > 0); } } break; case WAIT_OBJECT_0+1: { // Set the continue bit to false, so we'll exit fContinue = false; } break; default: { // Something went wrong //LOG( "Error while calling WaitForMultipleObjects" ); } break; } } while (fContinue); MessageBox(0,L"kka kk!!!",L"kka ga",0); return 0; } CPort::CPort(wchar_t *portName) { // Attempt to open the serial port (COM2) //lLastError = serial.Open(_T(portName),0,0,true); lLastError = serial.Open(portName,0,0,true); if (lLastError != ERROR_SUCCESS) { //LOG( "Unable to open COM-port" ); } // Setup the serial port (115200,8N1, which is the default setting) lLastError = serial.Setup(CSerial::EBaud115200,CSerial::EData8,CSerial::EParNone,CSerial::EStop1); if (lLastError != ERROR_SUCCESS) { //LOG( "Unable to set COM-port setting" ); } // Register only for the receive event lLastError = serial.SetMask(CSerial::EEventBreak | CSerial::EEventCTS | CSerial::EEventDSR | CSerial::EEventError | CSerial::EEventRing | CSerial::EEventRLSD | CSerial::EEventRecv); if (lLastError != ERROR_SUCCESS) { //LOG( "Unable to set COM-port event mask" ); } // Use 'non-blocking' reads, because we don't know how many bytes // will be received. This is normally the most convenient mode // (and also the default mode for reading data). lLastError = serial.SetupReadTimeouts(CSerial::EReadTimeoutNonblocking); if (lLastError != ERROR_SUCCESS) { //LOG( "Unable to set COM-port read timeout" ); } // Create a handle for the overlapped operations hevtOverlapped = ::CreateEvent(0,TRUE,FALSE,0);; if (hevtOverlapped == 0) { //LOG( "Unable to create manual-reset event for overlapped I/O" ); } // Setup the overlapped structure ov.hEvent = hevtOverlapped; // Open the "STOP" handle hevtStop = ::CreateEvent(0,TRUE,FALSE,_T("Overlapped_Stop_Event")); if (hevtStop == 0) { //LOG( "Unable to create manual-reset event for stop event" ); } HandleOfThread = CreateThread( NULL, 0, ThreadHandler, 0, 0, NULL); } CPort::~CPort() { //fContinue = false; //CloseHandle( HandleOfThread ); //serial.Close(); } void CPort::Close() { fContinue = false; CloseHandle( HandleOfThread ); serial.Close(); } wchar_t *CPort::GetData() { return data; } bool CPort::getCTS() { return serial.GetCTS(); } bool CPort::getDSR() { return serial.GetDSR(); } bool CPort::getRing() { return serial.GetRing(); } Unreal Script Code: class MyPlayerController extends GamePlayerController DLLBind(SerialPortDLL); dllimport final function Open(string portName); dllimport final function Close(); dllimport final function string GetData();

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  • Elegance, thy Name is jQuery

    - by SGWellens
    So, I'm browsing though some questions over on the Stack Overflow website and I found a good jQuery question just a few minutes old. Here is a link to it. It was a tough question; I knew that by answering it, I could learn new stuff and reinforce what I already knew: Reading is good, doing is better. Maybe I could help someone in the process too. I cut and pasted the HTML from the question into my Visual Studio IDE and went back to Stack Overflow to reread the question. Dang, someone had already answered it! And it was a great answer. I never even had a chance to start analyzing the issue. Now I know what a one-legged man feels like in an ass-kicking contest. Nevertheless, since the question and answer were so interesting, I decided to dissect them and learn as much as possible. The HTML consisted of some divs separated by h3 headings.  Note the elements are laid out sequentially with no programmatic grouping: <h3 class="heading">Heading 1</h3> <div>Content</div> <div>More content</div> <div>Even more content</div><h3 class="heading">Heading 2</h3> <div>some content</div> <div>some more content</div><h3 class="heading">Heading 3</h3> <div>other content</div></form></body>  The requirement was to wrap a div around each h3 heading and the subsequent divs grouping them into sections. Why? I don't know, I suppose if you screen-scrapped some HTML from another site, you might want to reformat it before displaying it on your own. Anyways… Here is the marvelously, succinct posted answer: $('.heading').each(function(){ $(this).nextUntil('.heading').andSelf().wrapAll('<div class="section">');}); I was familiar with all the parts except for nextUntil and andSelf. But, I'll analyze the whole answer for completeness. I'll do this by rewriting the posted answer in a different style and adding a boat-load of comments: function Test(){ // $Sections is a jQuery object and it will contain three elements var $Sections = $('.heading'); // use each to iterate over each of the three elements $Sections.each(function () { // $this is a jquery object containing the current element // being iterated var $this = $(this); // nextUntil gets the following sibling elements until it reaches // an element with the CSS class 'heading' // andSelf adds in the source element (this) to the collection $this = $this.nextUntil('.heading').andSelf(); // wrap the elements with a div $this.wrapAll('<div class="section" >'); });}  The code here doesn't look nearly as concise and elegant as the original answer. However, unless you and your staff are jQuery masters, during development it really helps to work through algorithms step by step. You can step through this code in the debugger and examine the jQuery objects to make sure one step is working before proceeding on to the next. It's much easier to debug and troubleshoot when each logical coding step is a separate line. Note: You may think the original code runs much faster than this version. However, the time difference is trivial: Not enough to worry about: Less than 1 millisecond (tested in IE and FF). Note: You may want to jam everything into one line because it results in less traffic being sent to the client. That is true. However, most Internet servers now compress HTML and JavaScript by stripping out comments and white space (go to Bing or Google and view the source). This feature should be enabled on your server: Let the server compress your code, you don't need to do it. Free Career Advice: Creating maintainable code is Job One—Maximum Priority—The Prime Directive. If you find yourself suddenly transferred to customer support, it may be that the code you are writing is not as readable as it could be and not as readable as it should be. Moving on… I created a CSS class to see the results: .section{ background-color: yellow; border: 2px solid black; margin: 5px;} Here is the rendered output before:   …and after the jQuery code runs.   Pretty Cool! But, while playing with this code, the logic of nextUntil began to bother me: What happens in the last section? What stops elements from being collected since there are no more elements with the .heading class? The answer is nothing.  In this case it stopped because it was at the end of the page.  But what if there were additional HTML elements? I added an anchor tag and another div to the HTML: <h3 class="heading">Heading 1</h3> <div>Content</div> <div>More content</div> <div>Even more content</div><h3 class="heading">Heading 2</h3> <div>some content</div> <div>some more content</div><h3 class="heading">Heading 3</h3> <div>other content</div><a>this is a link</a><div>unrelated div</div> </form></body> The code as-is will include both the anchor and the unrelated div. This isn't what we want.   My first attempt to correct this used the filter parameter of the nextUntil function: nextUntil('.heading', 'div')  This will only collect div elements. But it merely skipped the anchor tag and it still collected the unrelated div:   The problem is we need a way to tell the nextUntil function when to stop. CSS selectors to the rescue: nextUntil('.heading, a')  This tells nextUntil to stop collecting sibling elements when it gets to an element with a .heading class OR when it gets to an anchor tag. In this case it solved the problem. FYI: The comma operator in a CSS selector allows multiple criteria.   Bingo! One final note, we could have broken the code down even more: We could have replaced the andSelf function here: $this = $this.nextUntil('.heading, a').andSelf(); With this: // get all the following siblings and then add the current item$this = $this.nextUntil('.heading, a');$this.add(this);  But in this case, the andSelf function reads real nice. In my opinion. Here's a link to a jsFiddle if you want to play with it. I hope someone finds this useful Steve Wellens CodeProject

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  • Developing Schema Compare for Oracle (Part 1)

    - by Simon Cooper
    SQL Compare is one of Red Gate's most successful SQL Server tools; it allows developers and DBAs to compare and synchronize the contents of their databases. Although similar tools exist for Oracle, they are quite noticeably lacking in the usability and stability that SQL Compare is known for in the SQL Server world. We could see a real need for a usable schema comparison tools for Oracle, and so the Schema Compare for Oracle project was born. Over the next few weeks, as we come up to release of v1, I'll be doing a series of posts on the development of Schema Compare for Oracle. For the first post, I thought I would start with the main pitfalls that we stumbled across when developing the product, especially from a SQL Server background. 1. Schemas and Databases The most obvious difference is that the concept of a 'database' is quite different between Oracle and SQL Server. On SQL Server, one server instance has multiple databases, each with separate schemas. There is typically little communication between separate databases, and most databases are no more than about 1000-2000 objects. This means SQL Compare can register an entire database in a reasonable amount of time, and cross-database dependencies probably won't be an issue. It is a quite different scene under Oracle, however. The terms 'database' and 'instance' are used interchangeably, (although technically 'database' refers to the datafiles on disk, and 'instance' the running Oracle process that reads & writes to the database), and a database is a single conceptual entity. This immediately presents problems, as it is infeasible to register an entire database as we do in SQL Compare; in my Oracle install, using the standard recommended options, there are 63975 system objects. If we tried to register all those, not only would it take hours, but the client would probably run out of memory before we finished. As a result, we had to allow people to specify what schemas they wanted to register. This decision had quite a few knock-on effects for the design, which I will cover in a future post. 2. Connecting to Oracle The next obvious difference is in actually connecting to Oracle – in SQL Server, you can specify a server and database, and off you go. On Oracle things are slightly more complicated. SIDs, Service Names, and TNS A database (the files on disk) must have a unique identifier for the databases on the system, called the SID. It also has a global database name, which consists of a name (which doesn't have to match the SID) and a domain. Alternatively, you can identify a database using a service name, which normally has a 1-to-1 relationship with instances, but may not if, for example, using RAC (Real Application Clusters) for redundancy and failover. You specify the computer and instance you want to connect to using TNS (Transparent Network Substrate). The user-visible parts are a config file (tnsnames.ora) on the client machine that specifies how to connect to an instance. For example, the entry for one of my test instances is: SC_11GDB1 = (DESCRIPTION = (ADDRESS_LIST = (ADDRESS = (PROTOCOL = TCP)(HOST = simonctest)(PORT = 1521)) ) (CONNECT_DATA = (SID = 11gR1db1) ) ) This gives the hostname, port, and SID of the instance I want to connect to, and associates it with a name (SC_11GDB1). The tnsnames syntax also allows you to specify failover, multiple descriptions and address lists, and client load balancing. You can then specify this TNS identifier as the data source in a connection string. Although using ODP.NET (the .NET dlls provided by Oracle) was fine for internal prototype builds, once we released the EAP we discovered that this simply wasn't an acceptable solution for installs on other people's machines. Due to .NET assembly strong naming, users had to have installed on their machines the exact same version of the ODP.NET dlls as we had on our build server. We couldn't ship the ODP.NET dlls with our installer as the Oracle license agreement prohibited this, and we didn't want to force users to install another Oracle client just so they can run our program. To be able to list the TNS entries in the connection dialog, we also had to locate and parse the tnsnames.ora file, which was complicated by users with several Oracle client installs and intricate TNS entries. After much swearing at our computers, we eventually decided to use a third party Oracle connection library from Devart that we could ship with our program; this could use whatever client version was installed, parse the TNS entries for us, and also had the nice feature of being able to connect to an Oracle server without having any client installed at all. Unfortunately, their current license agreement prevents us from shipping an Oracle SDK, but that's a bridge we'll cross when we get to it. 3. Running synchronization scripts The most important difference is that in Oracle, DDL is non-transactional; you cannot rollback DDL statements like you can on SQL Server. Although we considered various solutions to this, including using the flashback archive or recycle bin, or generating an undo script, no reliable method of completely undoing a half-executed sync script has yet been found; so in this case we simply have to trust that the DBA or developer will check and verify the script before running it. However, before we got to that stage, we had to get the scripts to run in the first place... To run a synchronization script from SQL Compare we essentially pass the script over to the SqlCommand.ExecuteNonQuery method. However, when we tried to do the same for an OracleConnection we got a very strange error – 'ORA-00911: invalid character', even when running the most basic CREATE TABLE command. After much hair-pulling and Googling, we discovered that Oracle has got some very strange behaviour with semicolons at the end of statements. To understand what's going on, we need to take a quick foray into SQL and PL/SQL. PL/SQL is not T-SQL In SQL Server, T-SQL is the language used to interface with the database. It has DDL, DML, control flow, and many other nice features (like Turing-completeness) that you can mix and match in the same script. In Oracle, DDL SQL and PL/SQL are two completely separate languages, with different syntax, different datatypes and different execution engines within the instance. Oracle SQL is much more like 'pure' ANSI SQL, with no state, no control flow, and only the basic DML commands. PL/SQL is the Turing-complete language, but can only do DML and DCL (i.e. BEGIN TRANSATION commands). Any DDL or SQL commands that aren't recognised by the PL/SQL engine have to be passed back to the SQL engine via an EXECUTE IMMEDIATE command. In PL/SQL, a semicolons is a valid token used to delimit the end of a statement. In SQL, a semicolon is not a valid token (even though the Oracle documentation gives them at the end of the syntax diagrams) . When you execute the command CREATE TABLE table1 (COL1 NUMBER); in SQL*Plus the semicolon on the end is a command to SQL*Plus to execute the preceding statement on the server; it strips off the semicolon before passing it on. SQL Developer does a similar thing. When executing a PL/SQL block, however, the syntax is like so: BEGIN INSERT INTO table1 VALUES (1); INSERT INTO table1 VALUES (2); END; / In this case, the semicolon is accepted by the PL/SQL engine as a statement delimiter, and instead the / is the command to SQL*Plus to execute the current block. This explains the ORA-00911 error we got when trying to run the CREATE TABLE command – the server is complaining about the semicolon on the end. This also means that there is no SQL syntax to execute more than one DDL command in the same OracleCommand. Therefore, we would have to do a round-trip to the server for every command we want to execute. Obviously, this would cause lots of network traffic and be very slow on slow or congested networks. Our first attempt at a solution was to wrap every SQL statement (without semicolon) inside an EXECUTE IMMEDIATE command in a PL/SQL block and pass that to the server to execute. One downside of this solution is that we get no feedback as to how the script execution is going; we're currently evaluating better solutions to this thorny issue. Next up: Dependencies; how we solved the problem of being unable to register the entire database, and the knock-on effects to the whole product.

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  • krb5-multidev, libk5crypto3, libk5crypto3:i386 package dependency

    - by TDalton
    Using Ubuntu 12.04 Asus U43F I can no longer update, install or remove packages via software center because of package dependency errors. sudo apt-get install -f reads the following: Reading package lists... Done Building dependency tree Reading state information... Done Correcting dependencies... Done The following packages were automatically installed and are no longer required: libunity6 libqapt-runtime libboost-program-options1.46.1 akonadi-backend-mysql libqapt1 shared-desktop-ontologies libntrack0 ntrack-module-libnl-0 libntrack-qt4-1 Use 'apt-get autoremove' to remove them. The following extra packages will be installed: krb5-multidev libk5crypto3:i386 libkrb5-dev Suggested packages: krb5-doc krb5-doc:i386 krb5-user:i386 The following packages will be upgraded: krb5-multidev libk5crypto3:i386 libkrb5-dev 3 upgraded, 0 newly installed, 0 to remove and 325 not upgraded. 11 not fully installed or removed. Need to get 0 B/213 kB of archives. After this operation, 0 B of additional disk space will be used. Do you want to continue [Y/n]? y dpkg: error processing libk5crypto3 (--configure): libk5crypto3:amd64 1.10+dfsg~beta1-2ubuntu0.3 cannot be configured because libk5crypto3:i386 is in a different version (1.10+dfsg~beta1-2ubuntu0.1) dpkg: error processing libk5crypto3:i386 (--configure): libk5crypto3:i386 1.10+dfsg~beta1-2ubuntu0.1 cannot be configured because libk5crypto3:amd64 is in a different version (1.10+dfsg~beta1-2ubuntu0.3) dpkg: dependency problems prevent configuration of libkrb5-3: libkrb5-3 depends on libk5crypto3 (>= 1.9+dfsg~beta1); however:No apport report written because MaxReports is reached already Package libk5crypto3 is not configured yet. dpkg: error processing libkrb5-3 (--configure): dependency problems - leaving unconfigured dpkg: dependency problems prevent configuration of libgssapi-krb5-2: libgssapi-krb5-2 depends on libk5crypto3 (>= 1.8+dfsg); however: Package libk5crypto3 is not configured yet. libgssapi-krb5-2 depends on libkrb5-3 (= 1.10+dfsg~beta1-2ubuntu0.3); however: Package libkrb5-3 is not configured yet. dpkg: error processing libgssapi-krb5-2 (--configure): dependency problems - leaving unconfigured dpkg: dependency problems prevent configuration of libgssrpc4: libgssrpc4 depends on libgssapi-krb5-2 (>= 1.10+dfsg~); however: Package libgssapi-krb5-2 is not configured yet. dpkg: error processing libgssrpc4 (--configure): dependency problems - leaving unconfigured No apport report written because MaxReports is reached already No apport report written because MaxReports is reached already No apport report written because MaxReports is reached already dpkg: dependency problems prevent configuration of libkadm5srv-mit8: libkadm5srv-mit8 depends on libgssapi-krb5-2 (>= 1.6.dfsg.2); however: Package libgssapi-krb5-2 is not configured yet. libkadm5srv-mit8 depends on libgssrpc4 (>= 1.6.dfsg.2); however: Package libgssrpc4 is not configured yet. libkadm5srv-mit8 depends on libk5crypto3 (>= 1.6.dfsg.2); however: Package libk5crypto3 is not configured yet. libkadm5srv-mit8 depends on libkrb5-3 (>= 1.9+dfsg~beta1); however: Package libkrb5-3 is not configured yet. dpkg: error processing libkadm5srv-mit8 (--configure): dependency problems - leaving unconfigured dpkg: dependency problems prevent configuration of libkadm5clnt-mit8: libkadm5clnt-mit8 depends on libgssapi-krb5-2 (>= 1.10+dfsg~); however: Package libgssapi-krb5-2 is not configured yet. libkadm5clnt-mit8 depends on libgssrpc4 (>= 1.6.dfsg.2); however: Package libgssrpc4 is not configured yet.No apport report written because MaxReports is reached already libkadm5clnt-mit8 depends on libk5crypto3 (>= 1.6.dfsg.2); however: Package libk5crypto3 is not configured yet. libkadm5clnt-mit8 depends on libkrb5-3 (>= 1.8+dfsg); however: Package libkrb5-3 is not configured yet. dpkg: error processing libkadm5clnt-mit8 (--configure): dependency problems - leaving unconfigured No apport report written because MaxReports is reached already dpkg: dependency problems prevent configuration of krb5-multidev: krb5-multidev depends on libkrb5-3 (= 1.10+dfsg~beta1-2ubuntu0.2); however: Version of libkrb5-3 on system is 1.10+dfsg~beta1-2ubuntu0.3. krb5-multidev depends on libk5crypto3 (= 1.10+dfsg~beta1-2ubuntu0.2); however: Version of libk5crypto3 on system is 1.10+dfsg~beta1-2ubuntu0.3. krb5-multidev depends on libgssapi-krb5-2 (= 1.10+dfsg~beta1-2ubuntu0.2); however: Version of libgssapi-krb5-2 on system is 1.10+dfsg~beta1-2ubuntu0.3. krb5-multidev depends on libgssrpc4 (= 1.10+dfsg~beta1-2ubuntu0.2); however: Version of libgssrpc4 on system is 1.10+dfsg~beta1-2ubuntu0.3. krb5-multidev depends on libkadm5srv-mit8 (= 1.10+dfsg~beta1-2ubuntu0.2); however: Version of libkadm5srv-mit8 on system is 1.10+dfsg~beta1-2ubuntu0.3. krb5-multidev depends on libkadm5clnt-mit8 (= 1.10+dfsg~beta1-2ubuntu0.2); however: Version of libkadm5clnt-mit8 on system is 1.10+dfsg~beta1-2ubuntu0.3. dpkg: error processing krb5-multidev (--configure): dependency problems - leaving unconfigured No apport report written because MaxReports is reached already dpkg: dependency problems prevent configuration of libkrb5-dev: libkrb5-dev depends on krb5-multidev (= 1.10+dfsg~beta1-2ubuntu0.2); however: Package krb5-multidev is not configured yet. dpkg: error processing libkrb5-dev (--configure): dependency problems - leaving unconfigured No apport report written because MaxReports is reached already dpkg: dependency problems prevent configuration of libkrb5-3:i386: libkrb5-3:i386 depends on libk5crypto3 (>= 1.9+dfsg~beta1); however: Package libk5crypto3:i386 is not configured yet. dpkg: error processing libkrb5-3:i386 (--configure): dependency problems - leaving unconfigured dpkg: dependency problems prevent configuration of libgssapi-krb5-2:i386: libgssapi-krb5-2:i386 depends on libk5crypto3 (>= 1.8+dfsg); however: Package libk5crypto3:i386 is not configured yet. libgssapi-krb5-2:i386 depends on libkrb5-3 (= 1.10+dfsg~beta1-2ubuntu0.3); however: Package libkrb5-3:i386 is not configured yet. dpkg: error processing libgssapi-krb5-2:i386 (--configure): dependency problems - leaving unconfigured Errors were encountered while processing: libk5crypto3 libk5crypto3:i386 libkrb5-3 libgssapi-krb5-2 libgssrpc4 libkadm5srv-mit8 libkadm5clnt-mit8 krb5-multidev libkrb5-dev libkrb5-3:i386 libgssapi-krb5-2:i386 E: Sub-process /usr/bin/dpkg returned an error code (1) I have tried to fix the broken dependencies via synaptic package manager, but it returns with an error: E: libk5crypto3: libk5crypto3:amd64 1.10+dfsg~beta1-2ubuntu0.3 cannot be configured because libk5crypto3 E: libkrb5-3: dependency problems - leaving unconfigured E: libgssapi-krb5-2: dependency problems - leaving unconfigured E: libgssrpc4: dependency problems - leaving unconfigured E: libkadm5srv-mit8: dependency problems - leaving unconfigured E: libkadm5clnt-mit8: dependency problems - leaving unconfigured E: krb5-multidev: dependency problems - leaving unconfigured E: libkrb5-dev: dependency problems - leaving unconfigured I haven't gotten help from ubuntuforums.org on this issue. Please help, obi-wan

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  • postfix relaying all mail through office365 problems

    - by amrith
    This is a rather long question with a long list of things tried and travails so please bear with me. The summary is this. I am able to relay email from ubuntu through office365 using postfix; the configuration works. It only works as one of the users; more specifically the user who authenticates against office365 is the only valid "from" More details follow. I have a machine in Amazon's cloud on which I run a bunch of jobs and would like to have statuses mailed over to me. I use office365 at work so I want to relay mail through office365. I'm most familiar with postfix so I used that as the MTA. Configuration is ubuntu 12.04LTS; I've installed postfix and mail-utils. For this example, let me say my company is "company.com" and the machine in question (through an elastic IP and a DNS entry) is called "plaything.company.com". hostname is set to "plaything.company.com", so is /etc/mailname On plaything, I have the following users registered alpha, bravo, and charlie. I have the following configuration files. alias_database = hash:/etc/aliases alias_maps = hash:/etc/aliases append_dot_mydomain = no biff = no config_directory = /etc/postfix inet_interfaces = all inet_protocols = ipv4 mailbox_size_limit = 0 mydestination = plaything.company.com, localhost.company.com, , localhost myhostname = plaything.company.com mynetworks = 127.0.0.0/8 [::ffff:127.0.0.0]/104 [::1]/128 myorigin = /etc/mailname readme_directory = no recipient_delimiter = + relayhost = [smtp.office365.com]:587 sender_canonical_maps = hash:/etc/postfix/sender_canonical smtp_sasl_auth_enable = yes smtp_sasl_password_maps = hash:/etc/postfix/sasl_passwd smtp_sasl_security_options = noanonymous smtp_sasl_tls_security_options = noanonymous smtp_tls_CAfile = /etc/ssl/certs/ca-certificates.crt smtp_tls_session_cache_database = btree:${data_directory}/smtp_scache smtp_use_tls = yes smtpd_banner = $myhostname ESMTP $mail_name (Ubuntu) smtpd_tls_cert_file = /etc/ssl/certs/ssl-cert-snakeoil.pem smtpd_tls_key_file = /etc/ssl/private/ssl-cert-snakeoil.key smtpd_tls_session_cache_database = btree:${data_directory}/smtpd_scache smtpd_use_tls = yes As the machine is called plaything.company.com I went through the exercise of registering all the appropriate DNS entries to make office365 recognize that I owned plaything.company.com and allowed me to create a user called [email protected] in office365. In office365, I setup [email protected] as having another email address of [email protected]. Then, I made the following sender_canonical [email protected] [email protected] I created a sasl_passwd file that reads: smtp.office365.com [email protected]:123456password123456 let's just say that the password for [email protected] is 1234...456 With all this setup, login as alpha and mail [email protected] Cc: Subject: test test and the whole thing works wonderfully. email gets sent off by postfix, TLS works like a champ, authenticates as daemon@... and [email protected] in Office365 gets an email message. The issue comes up when logged in as bravo to the machine. sender is [email protected] and office365 says: status=bounced (host smtp.office365.com[132.245.12.25] said: 550 5.7.1 Client does not have permissions to send as this sender (in reply to end of DATA command)) this is because I'm trying to send mail as bravo@... and authenticating with office365 as daemon@.... The reason it works with alpha@... is because in office365, I setup [email protected] as having another email address of [email protected]. In Postfix Relay to Office365, Miles Erickson answers the question thusly: Don't send mail to Office365 as a user from your Office365-hosted e-mail domain. Use a subdomain instead, e.g. [email protected] instead of [email protected]. It wouldn't hurt to set up an SPF record for services.mydomain.com or whatever you decide to use. Don't authenticate against mail.messaging.microsoft.com as an Office365 user. Just connect on port 25 and deliver the mail to your domain as any foreign SMTP agent would do. OK, I've done #1, I have those records on DNS but for the most part they are not relevant once Office365 recognizes that I own the domain. Here are those records: CNAME records: - msoid.plaything.company.com - autodiscover.plaything.company.com MX record: - plaything.company.com (plaything-company-com.mail.protection.outlook.com) TXT record: - plaything.company.com (v=spf1 include:spf.protection.outlook.com -all) I've tried #2 but no matter what I do, office365 just blows away the connection with "not authenticated". I can try even a simple telnet to port 25 and attempt to send and it doesn't work. 250 BY2PR01CA007.outlook.office365.com Hello [54.221.245.236] 530 5.7.1 Client was not authenticated Connection closed by foreign host. Is there someone out there who has this kind of a configuration working where multiple users on a linux machine are able to relay mail using postfix through office365? There has to be someone out there doing this who can tell me what is wrong with my setup ...

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