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  • Navigation in a #WP7 application with MVVM Light

    - by Laurent Bugnion
    In MVVM applications, it can be a bit of a challenge to send instructions to the view (for example a page) from a viewmodel. Thankfully, we have good tools at our disposal to help with that. In his excellent series “MVVM Light Toolkit soup to nuts”, Jesse Liberty proposes one approach using the MVVM Light messaging infrastructure. While this works fine, I would like to show here another approach using what I call a “view service”, i.e. an abstracted service that is invoked from the viewmodel, and implemented on the view. Multiple kinds of view services In fact, I use view services quite often, and even started standardizing them for the Windows Phone 7 applications I work on. If there is interest, I will be happy to show other such view services, for example Animation services, responsible to start/stop animations on the view. Dialog service, in charge of displaying messages to the user and gathering feedback. Navigation service, in charge of navigating to a given page directly from the viewmodel. In this article, I will concentrate on the navigation service. The INavigationService interface In most WP7 apps, the navigation service is used in quite a straightforward way. We want to: Navigate to a given URI. Go back. Be notified when a navigation is taking place, and be able to cancel. The INavigationService interface is quite simple indeed: public interface INavigationService { event NavigatingCancelEventHandler Navigating; void NavigateTo(Uri pageUri); void GoBack(); } Obviously, this interface can be extended if necessary, but in most of the apps I worked on, I found that this covers my needs. The NavigationService class It is possible to nicely pack the navigation service into its own class. To do this, we need to remember that all the PhoneApplicationPage instances use the same instance of the navigation service, exposed through their NavigationService property. In fact, in a WP7 application, it is the main frame (RootFrame, of type PhoneApplicationFrame) that is responsible for this task. So, our implementation of the NavigationService class can leverage this. First the class will grab the PhoneApplicationFrame and store a reference to it. Also, it registers a handler for the Navigating event, and forwards the event to the listening viewmodels (if any). Then, the NavigateTo and the GoBack methods are implemented. They are quite simple, because they are in fact just a gateway to the PhoneApplicationFrame. The whole class is as follows: public class NavigationService : INavigationService { private PhoneApplicationFrame _mainFrame; public event NavigatingCancelEventHandler Navigating; public void NavigateTo(Uri pageUri) { if (EnsureMainFrame()) { _mainFrame.Navigate(pageUri); } } public void GoBack() { if (EnsureMainFrame() && _mainFrame.CanGoBack) { _mainFrame.GoBack(); } } private bool EnsureMainFrame() { if (_mainFrame != null) { return true; } _mainFrame = Application.Current.RootVisual as PhoneApplicationFrame; if (_mainFrame != null) { // Could be null if the app runs inside a design tool _mainFrame.Navigating += (s, e) => { if (Navigating != null) { Navigating(s, e); } }; return true; } return false; } } Exposing URIs I find that it is a good practice to expose each page’s URI as a constant. In MVVM Light applications, a good place to do that is the ViewModelLocator, which already acts like a central point of setup for the views and their viewmodels. Note that in some cases, it is necessary to expose the URL as a string, for instance when a query string needs to be passed to the view. So for example we could have: public static readonly Uri MainPageUri = new Uri("/MainPage.xaml", UriKind.Relative); public const string AnotherPageUrl = "/AnotherPage.xaml?param1={0}&param2={1}"; Creating and using the NavigationService Normally, we only need one instance of the NavigationService class. In cases where you use an IOC container, it is easy to simply register a singleton instance. For example, I am using a modified version of a super simple IOC container, and so I can register the navigation service as follows: SimpleIoc.Register<INavigationService, NavigationService>(); Then, it can be resolved where needed with: SimpleIoc.Resolve<INavigationService>(); Or (more frequently), I simply declare a parameter on the viewmodel constructor of type INavigationService and let the IOC container do its magic and inject the instance of the NavigationService when the viewmodel is created. On supported platforms (for example Silverlight 4), it is also possible to use MEF. Or, of course, we can simply instantiate the NavigationService in the ViewModelLocator, and pass this instance as a parameter of the viewmodels’ constructor, injected as a property, etc… Once the instance has been passed to the viewmodel, it can be used, for example with: NavigationService.NavigateTo(ViewModelLocator.ComparisonPageUri); Testing Thanks to the INavigationService interface, navigation can be mocked and tested when the viewmodel is put under unit test. Simply implement and inject a mock class, and assert that the methods are called as they should by the viewmodel. Conclusion As usual, there are multiple ways to code a solution answering your needs. I find that view services are a really neat way to delegate view-specific responsibilities such as animation, dialogs and of course navigation to other classes through an abstracted interface. In some cases, such as the NavigationService class exposed here, it is even possible to standardize the implementation and pack it in a class library for reuse. I hope that this sample is useful! Happy coding. Laurent   Laurent Bugnion (GalaSoft) Subscribe | Twitter | Facebook | Flickr | LinkedIn

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  • Battery life starts at 2:30 hrs (99%), but less than 1 minute later is only 1:30 hrs (99%)

    - by zondu
    After searching this and other forums, I haven't seen this same issue listed anywhere for Ubuntu 12. Prior to installing Ubuntu 12.10, my Netbook (Acer AspireOne D250, SATA HDD) was consistently getting 2:30-3 hrs battery life under Windows XP Home, SP3. However, immediately after installing Ubuntu 12.10, the battery life starts out at 2:30 hrs (99%), but less than 1 minute later suddenly drops to 1:30 hrs (99%), which seems very odd. It could be a complete coincidence that the battery is suddenly flaky at the exact same moment that Ubuntu 12.10 was installed, but that doesn't seem likely. I'm a newbie to Ubuntu, so I don't have much experience tweaking/trouble-shooting yet. Here's what I've tried so far: enabled laptop mode (sudo su, then echo 5 /proc/sys/vm/laptop_mode) and checked that it is running when the A/C adapter is unplugged, but it doesn't seem to have made any noticeable difference in battery life, installed Jupiter, but it didn't work and messed up the system, so I had to uninstall it, disabled bluetooth (wifi is still on b/c it is necessary), set the screen to lowest brightness, etc., run through at least 1 full power cycle (running until the netbook shut itself off due to critical battery) and have been using it normally (sometimes plugged in, often unplugged until the battery gets very low) for a week since installing Ubuntu 12.10. installed powertop, but have no idea how to interpret its results. Here are the results of acpi -b: w/ A/C adapter: Battery 0: Full, 100% immediately after unplugging: Battery 0: Discharging, 99%, 02:30:20 remaining 1 minute after unplugging: Battery 0: Discharging, 99%, 01:37:49 remaining 2-3 minutes after unplugging: Battery 0: Discharging, 95%, 01:33:01 remaining 10 minutes after unplugging: Battery 0: Discharging, 85%, 01:13:38 remaining Results of cat /sys/class/power_supply/BAT0/uevent: w/ A/C adapter: POWER_SUPPLY_NAME=BAT0 POWER_SUPPLY_STATUS=Full POWER_SUPPLY_PRESENT=1 POWER_SUPPLY_TECHNOLOGY=Li-ion POWER_SUPPLY_CYCLE_COUNT=0 POWER_SUPPLY_VOLTAGE_MIN_DESIGN=10800000 POWER_SUPPLY_VOLTAGE_NOW=12136000 POWER_SUPPLY_CURRENT_NOW=773000 POWER_SUPPLY_CHARGE_FULL_DESIGN=4500000 POWER_SUPPLY_CHARGE_FULL=1956000 POWER_SUPPLY_CHARGE_NOW=1956000 POWER_SUPPLY_MODEL_NAME=UM08B32 POWER_SUPPLY_MANUFACTURER=SANYO POWER_SUPPLY_SERIAL_NUMBER= immediately after unplugging: POWER_SUPPLY_NAME=BAT0 POWER_SUPPLY_STATUS=Discharging POWER_SUPPLY_PRESENT=1 POWER_SUPPLY_TECHNOLOGY=Li-ion POWER_SUPPLY_CYCLE_COUNT=0 POWER_SUPPLY_VOLTAGE_MIN_DESIGN=10800000 POWER_SUPPLY_VOLTAGE_NOW=11886000 POWER_SUPPLY_CURRENT_NOW=773000 POWER_SUPPLY_CHARGE_FULL_DESIGN=4500000 POWER_SUPPLY_CHARGE_FULL=1956000 POWER_SUPPLY_CHARGE_NOW=1937000 POWER_SUPPLY_MODEL_NAME=UM08B32 POWER_SUPPLY_MANUFACTURER=SANYO POWER_SUPPLY_SERIAL_NUMBER= 1 minute later: POWER_SUPPLY_NAME=BAT0 POWER_SUPPLY_STATUS=Discharging POWER_SUPPLY_PRESENT=1 POWER_SUPPLY_TECHNOLOGY=Li-ion POWER_SUPPLY_CYCLE_COUNT=0 POWER_SUPPLY_VOLTAGE_MIN_DESIGN=10800000 POWER_SUPPLY_VOLTAGE_NOW=11728000 POWER_SUPPLY_CURRENT_NOW=1174000 POWER_SUPPLY_CHARGE_FULL_DESIGN=4500000 POWER_SUPPLY_CHARGE_FULL=1956000 POWER_SUPPLY_CHARGE_NOW=1937000 POWER_SUPPLY_MODEL_NAME=UM08B32 POWER_SUPPLY_MANUFACTURER=SANYO POWER_SUPPLY_SERIAL_NUMBER= 2-3 minutes later: POWER_SUPPLY_NAME=BAT0 POWER_SUPPLY_STATUS=Discharging POWER_SUPPLY_PRESENT=1 POWER_SUPPLY_TECHNOLOGY=Li-ion POWER_SUPPLY_CYCLE_COUNT=0 POWER_SUPPLY_VOLTAGE_MIN_DESIGN=10800000 POWER_SUPPLY_VOLTAGE_NOW=11583000 POWER_SUPPLY_CURRENT_NOW=1209000 POWER_SUPPLY_CHARGE_FULL_DESIGN=4500000 POWER_SUPPLY_CHARGE_FULL=1956000 POWER_SUPPLY_CHARGE_NOW=1878000 POWER_SUPPLY_MODEL_NAME=UM08B32 POWER_SUPPLY_MANUFACTURER=SANYO POWER_SUPPLY_SERIAL_NUMBER= 10 minutes later: POWER_SUPPLY_NAME=BAT0 POWER_SUPPLY_STATUS=Discharging POWER_SUPPLY_PRESENT=1 POWER_SUPPLY_TECHNOLOGY=Li-ion POWER_SUPPLY_CYCLE_COUNT=0 POWER_SUPPLY_VOLTAGE_MIN_DESIGN=10800000 POWER_SUPPLY_VOLTAGE_NOW=11230000 POWER_SUPPLY_CURRENT_NOW=1239000 POWER_SUPPLY_CHARGE_FULL_DESIGN=4500000 POWER_SUPPLY_CHARGE_FULL=1956000 POWER_SUPPLY_CHARGE_NOW=1644000 POWER_SUPPLY_MODEL_NAME=UM08B32 POWER_SUPPLY_MANUFACTURER=SANYO POWER_SUPPLY_SERIAL_NUMBER= Results of upower -i /org/freedesktop/UPower/devices/battery_BAT0: w/ A/C adapter: native-path: /sys/devices/LNXSYSTM:00/device:00/PNP0A08:00/device:02/PNP0C0A:00/power_supply/BAT0 vendor: SANYO model: UM08B32 power supply: yes updated: Tue Nov 27 15:24:58 2012 (823 seconds ago) has history: yes has statistics: yes battery present: yes rechargeable: yes state: fully-charged energy: 21.1248 Wh energy-empty: 0 Wh energy-full: 21.1248 Wh energy-full-design: 48.6 Wh energy-rate: 8.3484 W voltage: 12.173 V percentage: 100% capacity: 43.4667% technology: lithium-ion immediately after unplugging: native-path: /sys/devices/LNXSYSTM:00/device:00/PNP0A08:00/device:02/PNP0C0A:00/power_supply/BAT0 vendor: SANYO model: UM08B32 power supply: yes updated: Tue Nov 27 15:41:25 2012 (1 seconds ago) has history: yes has statistics: yes battery present: yes rechargeable: yes state: discharging energy: 20.9196 Wh energy-empty: 0 Wh energy-full: 21.1248 Wh energy-full-design: 48.6 Wh energy-rate: 8.3484 W voltage: 11.86 V time to empty: 2.5 hours percentage: 99.0286% capacity: 43.4667% technology: lithium-ion History (charge): 1354023683 99.029 discharging 1 minute later: native-path: /sys/devices/LNXSYSTM:00/device:00/PNP0A08:00/device:02/PNP0C0A:00/power_supply/BAT0 vendor: SANYO model: UM08B32 power supply: yes updated: Tue Nov 27 15:42:31 2012 (17 seconds ago) has history: yes has statistics: yes battery present: yes rechargeable: yes state: discharging energy: 20.9196 Wh energy-empty: 0 Wh energy-full: 21.1248 Wh energy-full-design: 48.6 Wh energy-rate: 13.5432 W voltage: 11.753 V time to empty: 1.5 hours percentage: 99.0286% capacity: 43.4667% technology: lithium-ion History (charge): 1354023683 99.029 discharging History (rate): 1354023751 13.543 discharging 2-3 minutes later: native-path: /sys/devices/LNXSYSTM:00/device:00/PNP0A08:00/device:02/PNP0C0A:00/power_supply/BAT0 vendor: SANYO model: UM08B32 power supply: yes updated: Tue Nov 27 15:45:06 2012 (20 seconds ago) has history: yes has statistics: yes battery present: yes rechargeable: yes state: discharging energy: 20.2824 Wh energy-empty: 0 Wh energy-full: 21.1248 Wh energy-full-design: 48.6 Wh energy-rate: 13.7484 W voltage: 11.545 V time to empty: 1.5 hours percentage: 96.0123% capacity: 43.4667% technology: lithium-ion History (charge): 1354023906 96.012 discharging 1354023844 97.035 discharging History (rate): 1354023906 13.748 discharging 1354023875 12.992 discharging 1354023844 13.284 discharging 10 minutes later: native-path: /sys/devices/LNXSYSTM:00/device:00/PNP0A08:00/device:02/PNP0C0A:00/power_supply/BAT0 vendor: SANYO model: UM08B32 power supply: yes updated: Tue Nov 27 15:54:24 2012 (28 seconds ago) has history: yes has statistics: yes battery present: yes rechargeable: yes state: discharging energy: 18.1764 Wh energy-empty: 0 Wh energy-full: 21.1248 Wh energy-full-design: 48.6 Wh energy-rate: 13.2948 W voltage: 11.268 V time to empty: 1.4 hours percentage: 86.0429% capacity: 43.4667% technology: lithium-ion History (charge): 1354024433 86.043 discharging History (rate): 1354024464 13.295 discharging 1354024433 13.662 discharging 1354024402 13.781 discharging I noticed that between #2 and #3 (0 and 1 minutes after unplugging), while the battery still reports 99% charge and drops from 2:30 hr to 1:30 hr, the energy usage goes from 8.34 W to 13.54 W and the current_now increases, but shouldn't it be using less energy in battery mode since the screen is much dimmer and it's in power saving mode? (or is that normal behavior?) It also seems to drain more quickly than what it predicts, especially with the 1-1.25 hour drop in the first minute of being unplugged, which seems odd. What really concerns me is that Ubuntu 12.10 may not be properly managing the battery (with the sudden change in charge/life from 2:30 to 1:30 or 1:15 within a minute of unplugging), and that a new battery may quickly die under Ubuntu 12.10. I'd greatly appreciate any advice/suggestions on what to do, and especially whether there's a way to get back the 1-1.5 hrs of battery life that were suddenly lost when changing from WinXp to Ubuntu 12.10. Thanks :)

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  • Migrating R Scripts from Development to Production

    - by Mark Hornick
    Normal 0 false false false EN-US X-NONE X-NONE MicrosoftInternetExplorer4 “How do I move my R scripts stored in one database instance to another? I have my development/test system and want to migrate to production.” Users of Oracle R Enterprise Embedded R Execution will often store their R scripts in the R Script Repository in Oracle Database, especially when using the ORE SQL API. From previous blog posts, you may recall that Embedded R Execution enables running R scripts managed by Oracle Database using both R and SQL interfaces. In ORE 1.3.1., the SQL API requires scripts to be stored in the database and referenced by name in SQL queries. The SQL API enables seamless integration with database-based applications and ease of production deployment. Loading R scripts in the repository Before talking about migration, we’ll first introduce how users store R scripts in Oracle Database. Users can add R scripts to the repository in R using the function ore.scriptCreate, or SQL using the function sys.rqScriptCreate. For the sample R script     id <- 1:10     plot(1:100,rnorm(100),pch=21,bg="red",cex =2)     data.frame(id=id, val=id / 100) users wrap this in a function and store it in the R Script Repository with a name. In R, this looks like ore.scriptCreate("RandomRedDots", function () { line-height: 115%; font-family: "Courier New";">     id <- 1:10     plot(1:100,rnorm(100),pch=21,bg="red",cex =2)     data.frame(id=id, val=id / 100)) }) In SQL, this looks like begin sys.rqScriptCreate('RandomRedDots',  'function(){     id <- 1:10     plot(1:100,rnorm(100),pch=21,bg="red",cex =2)     data.frame(id=id, val=id / 100)   }'); end; / The R function ore.scriptDrop and SQL function sys.rqScriptDrop can be used to drop these scripts as well. Note that the system will give an error if the script name already exists. Accessing R scripts once they’ve been loaded If you’re not using a source code control system, it is possible that your R scripts can be misplaced or files modified, making what is stored in Oracle Database to only or best copy of your R code. If you’ve loaded your R scripts to the database, it is straightforward to access these scripts from the database table SYS.RQ_SCRIPTS. For example, select * from sys.rq_scripts where name='myScriptName'; From R, scripts in the repository can be loaded into the R client engine using a function similar to the following: ore.scriptLoad <- function(name) { query <- paste("select script from sys.rq_scripts where name='",name,"'",sep="") str.f <- OREbase:::.ore.dbGetQuery(query) assign(name,eval(parse(text = str.f)),pos=1) } ore.scriptLoad("myFunctionName") This function is also useful if you want to load an existing R script from the repository into another R script in the repository – think modular coding style. Just include this function in the body of the other function and load the named script. Migrating R scripts from one database instance to another To move a set of functions from one system to another, the following script loads the functions from one R script repository into the client R engine, then connects to the target database and creates the scripts there with the same names. scriptNames <- OREbase:::.ore.dbGetQuery("select name from sys.rq_scripts where name not like 'RQG$%' and name not like 'RQ$%'")$NAME for(s in scriptNames) { cat(s,"\n") ore.scriptLoad(s) } ore.disconnect() ore.connect("rquser","orcl","localhost","rquser") for(s in scriptNames) { cat(s,"\n") ore.scriptDrop(s) ore.scriptCreate(s,get(s)) } Best Practice When naming R scripts, keep in mind that the name can be up to 128 characters. As such, consider organizing scripts in a directory structure manner. For example, if an organization has multiple groups or applications sharing the same database and there are multiple components, use “/” to facilitate the function organization: line-height: 115%;">ore.scriptCreate("/org1/app1/component1/myFuntion1", myFunction1) ore.scriptCreate("/org1/app1/component1/myFuntion2", myFunction2) ore.scriptCreate("/org1/app2/component2/myFuntion2", myFunction2) ore.scriptCreate("/org2/app2/component1/myFuntion3", myFunction3) ore.scriptCreate("/org3/app2/component1/myFuntion4", myFunction4) Users can then query for all functions using the path prefix when looking up functions. /* Style Definitions */ table.MsoNormalTable {mso-style-name:"Table Normal"; mso-tstyle-rowband-size:0; mso-tstyle-colband-size:0; mso-style-noshow:yes; mso-style-priority:99; mso-style-qformat:yes; mso-style-parent:""; mso-padding-alt:0in 5.4pt 0in 5.4pt; mso-para-margin:0in; mso-para-margin-bottom:.0001pt; mso-pagination:widow-orphan; font-size:11.0pt; font-family:"Calibri","sans-serif"; mso-ascii-font-family:Calibri; mso-ascii-theme-font:minor-latin; mso-fareast-font-family:"Times New Roman"; mso-fareast-theme-font:minor-fareast; mso-hansi-font-family:Calibri; mso-hansi-theme-font:minor-latin; mso-bidi-font-family:"Times New Roman"; mso-bidi-theme-font:minor-bidi;}

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  • SOA, Governance, and Drugs

    Why is IT governance important in service oriented architecture (SOA)? IT Governance provides a framework for making appropriate decisions based on company guidelines and accepted standards. This framework also outlines each stakeholder’s responsibilities and authority when making important architectural or design decisions. Furthermore, this framework of governance defines parameters and constraints that are used to give context and perspective when making decisions. The use of governance as it applies to SOA ensures that specific design principles and patterns are used when developing and maintaining services. When governance is consistently applied systems the following benefits are achieved according to Anne Thomas Manes in 2010. Governance makes sure that services conform to standard interface patterns, common data modeling practices, and promotes the incorporation of existing system functionality by building on top of other available services across a system. Governance defines development standards based on proven design principles and patterns that promote reuse and composition. Governance provides developers a set of proven design principles, standards and practices that promote the reduction in system based component dependencies.  By following these guidelines, individual components will be easier to maintain. For me personally, I am a fan of IT governance, and feel that it valuable part of any corporate IT department. However, depending on how it is implemented can really affect the value of using IT governance.  Companies need to find a way to ensure that governance does not become extreme in its policies and procedures. I know for me personally, I would really dislike working under a completely totalitarian or laissez-faire version of governance. Developers need to be able to be creative in their designs and too much governance can really impede the design process and prevent the most optimal design from being developed. On the other hand, with no governance enforced, no standards will be followed and accepted design patterns will be ignored. I have personally had to spend a lot of time working on this particular scenario and I have found that the concept of code reuse and composition is almost nonexistent.  Based on this, too much time and money is wasted on redeveloping existing aspects of an application that already exist within the system as a whole. I think moving forward we will see a staggered form of IT governance, regardless if it is for SOA or IT in general.  Depending on the size of a company and the size of its IT department,  I can see IT governance as a layered approach in that the top layer will be defined by enterprise architects that focus on abstract concepts pertaining to high level design, general  guidelines, acceptable best practices, and recommended design patterns.  The next layer will be defined by solution architects or department managers that further expand on abstracted guidelines defined by the enterprise architects. This layer will contain further definitions as to when various design patterns, coding standards, and best practices are to be applied based on the context of the solutions that are being developed by the department. The final layer will be defined by the system designer or a solutions architect assed to a project in that they will define what design patterns will be used in a solution, naming conventions, as well as outline how a system will function based on the best practices defined by the previous layers. This layered approach allows for IT departments to be flexible in that system designers have creative leeway in designing solutions to meet the needs of the business, but they must operate within the confines of the abstracted IT governance guidelines.  A real world example of this can be seen in the United States as it pertains to governance of the people in that the US government defines rules and regulations in the abstract and then the state governments take these guidelines and applies them based on the will of the people in each individual state. Furthermore, the county or city governments are the ones that actually enforce these rules based on how they are interpreted by local community.  To further define my example, the United States government defines that marijuana is illegal. Each individual state has the option to determine this regulation as it wishes in that the state of Florida determines that all uses of the drug are illegal, but the state of California legally allows the use of marijuana for medicinal purposes only. Based on these accepted practices each local government enforces these rules in that a police officer will arrest anyone in the state of Florida for having this drug on them if they walk down the street, but in California if a person has a medical prescription for the drug they will not get arrested.  REFERENCESThomas Manes, Anne. (2010). Understanding SOA Governance: http://www.soamag.com/I40/0610-2.php

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  • The design of a generic data synchronizer, or, an [object] that does [actions] with the aid of [helpers]

    - by acheong87
    I'd like to create a generic data-source "synchronizer," where data-source "types" may include MySQL databases, Google Spreadsheets documents, CSV files, among others. I've been trying to figure out how to structure this in terms of classes and interfaces, keeping in mind (what I've read about) composition vs. inheritance and is-a vs. has-a, but each route I go down seems to violate some principle. For simplicity, assume that all data-sources have a header-row-plus-data-rows format. For example, assume that the first rows of Google Spreadsheets documents and CSV files will have column headers, a.k.a. "fields" (to parallel database fields). Also, eventually, I would like to implement this in PHP, but avoiding language-specific discussion would probably be more productive. Here's an overview of what I've tried. Part 1/4: ISyncable class CMySQL implements ISyncable GetFields() // sql query, pdo statement, whatever AddFields() RemFields() ... _dbh class CGoogleSpreadsheets implements ISyncable GetFields() // zend gdata api AddFields() RemFields() ... _spreadsheetKey _worksheetId class CCsvFile implements ISyncable GetFields() // read from buffer AddFields() RemFields() ... _buffer interface ISyncable GetFields() AddFields($field1, $field2, ...) RemFields($field1, $field2, ...) ... CanAddFields() // maybe the spreadsheet is locked for write, or CanRemFields() // maybe no permission to alter a database table ... AddRow() ModRow() RemRow() ... Open() Close() ... First Question: Does it make sense to use an interface, as above? Part 2/4: CSyncer Next, the thing that does the syncing. class CSyncer __construct(ISyncable $A, ISyncable $B) Push() // sync A to B Pull() // sync B to A Sync() // Push() and Pull() only differ in direction; factor. // Sync()'s job is to make sure that the fields on each side // match, to add fields where appropriate and possible, to // account for different column-orderings, etc., and of // course, to add and remove rows as necessary to sync. ... _A _B Second Question: Does it make sense to define such a class, or am I treading dangerously close to the "Kingdom of Nouns"? Part 3/4: CTranslator? ITranslator? Now, here's where I actually get lost, assuming the above is passable. Sometimes, two ISyncables speak different "dialects." For example, believe it or not, Google Spreadsheets (accessed through the Google Data API "list feed") returns column headers lower-cased and stripped of all spaces and symbols! That is, sys_TIMESTAMP is systimestamp, as far as my code can tell. (Yes, I am aware that the "cell feed" does not strip the name so; however cell-by-cell manipulation is too slow for what I'm doing.) One can imagine other hypothetical examples. Perhaps even the data itself can be in different "dialects." But let's take it as given for now, and not argue this if possible. Third Question: How would you implement "translation"? Note: Taking all this as an exercise, I'm more interested in the "idealized" design, rather than the practical one. (God knows that shipped sailed when I began this project.) Part 4/4: Further Thought Here's my train of thought to demonstrate I've thunk, albeit unfruitfully: First, I thought, primitively, "I'll just modify CMySQL::GetFields() to lower-case and strip field names so they're compatible with Google Spreadsheets." But of course, then my class should really be called, CMySQLForGoogleSpreadsheets, and that can't be right. So, the thing which translates must exist outside of an ISyncable implementor. And surely it can't be right to make each translation a method in CSyncer. If it exists outside of both ISyncable and CSyncer, then what is it? (Is it even an "it"?) Is it an abstract class, i.e. abstract CTranslator? Is it an interface, since a translator only does, not has, i.e. interface ITranslator? Does it even require instantiation? e.g. If it's an ITranslator, then should its translation methods be static? (I learned what "late static binding" meant, today.) And, dear God, whatever it is, how should a CSyncer use it? Does it "have" it? Is it, "it"? Who am I? ...am I, "I"? I've attempted to break up the question into sub-questions, but essentially my question is singular: How does one implement an object A that conceptually "links" (has) two objects b1 and b2 that share a common interface B, where certain pairs of b1 and b2 require a helper, e.g. a translator, to be handled by A? Something tells me that I've overcomplicated this design, or violated a principle much higher up. Thank you all very much for your time and any advice you can provide.

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  • The Dreaded Startup Repair Loop on Win 7

    - by HighAltitudeCoder
    For most people, upgrading to Windows 7 has been a relatively painless process.  Not me.  I am in the unlucky 1% or less who had a somewhat less pleasant experience.  First, I cloned my entire onto a larger (and much faster) solid state hard drive, only experiencing minimal problems. Then, I bought the Retail version of Windows 7 Ultimate, took a deep breath and... oh yeah, I almost forgot - BACK UP THE COMPUTER.  The next morning I upgraded to Win 7 and everything seemed fine, until... I rebooted the system, the nice Windows 7 launch graphics come up, it's about to launch and AWWW, are you kidding me?!?!  Back to the BIOS splash screen?  Next comes the sequence of failure - attempt repair - unable to repair - do you want to wipe your hard drive decisions. Because I purchased the retail version, a number is provided where I could call Microsoft Tech support.  When I did, they instructed me to click "Install" from my installation CD, which did not work.  When I tried the "Upgrade" option, it reaches an impasse, telling you that yoiu have a newer version of Win 7, and thus cannot Upgrade.  If you choose "Install" you willl lose everything... files, programs, EVERYTHING.  Or at least this is what it tells you.  I was not willing to take the risk. To make things worse, I had installed a new antivirus software application before I realized my system was unstable (Trend Micro Titanium Internet Security), and this was causing additional problems. One interesting thing, and the only saving grace as it turns out, was that my system WOULD successfully reboot into the OS if I chose to restart it, rather than shut it down.  If I chose to shut down, I would have to go through the loop again until I was given the option to restart. As it turned out, I needed to update my BIOS.  I assumed that since I had updated my BIOS a long time ago to settings that were stable under Windows Vista Ultimate x64, I incorrectly expected Win 7 to adopt the same settings and didn't expect there to be any problems.  WRONG. My BIOS had a setting to halt the boot cycle if various kinds of errors were detected.  Windows Vista didn't care about this, but forget it under Windows 7.  I turned immediately corrected that BIOS setting.  Next, there were the two separate BIOS settings: enable USB mouse and enable USB keyboard.  The only sequence of events that would work were to start my reboot process over from stratch with a hard-wired non-usb keyboard and mouse.  Whent the system booted under these settings, it doesn't detect any errors due to either the mouse or keyboard, and actually booted for the first time in a long while (let me tell ya, that's an amazing experience after fiddling with settings for two entire weekends!) Next step: leave your old mouse and keyboard connected, but also connect your other two devices (mouse, keyboard) that use USB connections.  During the boot cycle, the operating system will not fail due to missing requirements during startup, and it will then pick up the new drivers necessary to use your new hardware. If you think you are in the clear here, you are wrong.  The next VERY IMPORTANT step is to remember to change your settings in the BIOS upon next startup.  Specifically, yoiu will need ot change your BIOS to enable USB mouse and enable USB keyboard input.  If you don't, Windows will detect an incompatibility upon the next startup, and you will be stuck once again in the endless cycle of reboot/Startup Repair/reboot/Startup Repair, without ever reaching a successful boot. Here's the thing - the BIOS and the drivers registered in Win 7 need to match.  If they don't, you're going to lose another weekend worrying and fiddling, all the while wondering if you've permanently damaged your hard drive beyond repair. (Sigh).  In the end, things worked out.  I must note that it is saddening to see how many posts there are out there that recommend just doing a clean install, as if it's the only option.  How many countless poor souls have lost their data, their backups, their pictures and videos, all for nothing other than the fact that the person giving advice just didn't know what to do at that point? My advice to you, try having a look at your BIOS settings first and making sure Win 7 can find your BIOS settings, and also disabling in your BIOS anything that might halt your system boot-up process if it encounters errors.

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  • Adventures in Windows 8: Solving activation errors

    - by Laurent Bugnion
    Note: I tagged this article with “MVVM” because I got a few support requests for MVVM Light regarding this exact issue. In fact it is a Windows 8 issue and has nothing to do with MVVM Light per se… Sometimes when you work on a Windows 8 app, you will get a very annoying issue when starting the app. In that case, the app doesn’t not even start past the Splash screen. Putting a breakpoint in App.xaml.cs doesn’t help because the app doesn’t even reach that point! So what exactly is happening? Well when a Windows 8 app starts, the system is performing a few check first. One of the checks, for instance, is to see if an app with the same package ID is already available. The package ID is a unique value set in the package manifest. In the Solution Explorer, double click on Package.appxmanifest. This opens the manifest in a special editor Click on the Packaging tab See the GUID under Package Name. This is the unique ID I am talking about. If there is a conflict (i.e. if an app is already installed with the exact same ID), Windows will warm the user that the app is already installed. However when you are in the process of developing an app, you install and uninstall the same app many many times (every time that you start in Visual Studio), and sometimes some issues arise, for instance failing to uninstall the app before starting the new instance of the same app. First step if you get such an error When the application fails to start past the splash screen, the first step is to identify what kind of error happened. In my experience the “already installed” is by far the most frequent (in fact I never had another such error), but it can be something else. An annoying thing is that the popup that shows the error is usually started below the Windows 8 app, and so you don’t even see it! This is especially true if you run this in the Simulator. In that case, do the following: Press on the Simulator’s home button, then press on the Desktop tile on the Start menu. The error popup should be shown on the desktop. If your applications runs on the Local machine, you also do the same and press the Windows button, and then from the Start menu press the Desktop tile. Deployment error in Studio Sometimes the same error causes Visual Studio to fail launching the application at all with a deployment error. This is a better case, because at least it is clear that there is an issue. In that case, write down the code that is shown in the Error window (for instance 0x80073D05 in the example below). Once you have the error code, go to the “Troubleshooting packaging, deployment, and query of Windows Store apps” page and look up the code in question. In my case, the error was “ERROR_DELETING_EXISTING_APPLICATIONDATA_STORE_FAILED”, “An error occurred while deleting the package's previously existing application data.” Solving the “ERROR_DELETING_EXISTING_APPLICATIONDATA_STORE_FAILED” issue Update: Before trying the below, you can also try the simple steps: Exit Visual Studio Go to the Start menu Locate your app’s tile. It should be visible in the Start menu directly, towards the far end on the right. Right click the tile and select Uninstall from the App Bar. Restart Visual Studio and try again. Sometimes it helps. If it doesn’t, then try the following: In order to solve the case where Windows, for any reason, fails to delete the existing application before starting the new instance, follow the steps: Open the Package.appxmanifest in Visual Studio Open the Packaging tab. Change the Package name. For tests you can just try to change the last character of the GUID, though I would recommend creating a brand new GUID. Press Start Type GUID Start the GUID Generator application Select Registry Format Press Copy. Paste the new GUID in place of the Package Name in Visual Studio Important: don’t forget to remove the curly brackets at the beginning and at the end of the newly pasted GUID. Then you just have to cross your fingers and start the application again… If it works, celebrate. if it doesn’t work… well at this point I am not sure so good luck with that ;) Happy coding, Laurent   Laurent Bugnion (GalaSoft) Subscribe | Twitter | Facebook | Flickr | LinkedIn

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  • javaf, problem...plz help someone...urgent [closed]

    - by innovative_aj
    i have made a word guessing game, when i click myButton to check if the guessed word is right or wrong, ball1 is moved into the "container" if its right, i want that when i click the button again and if the typed word is right, the 2nd ball should move into the container too... means one ball per correct answer...plz help me someone and provide me with the code that i can implement, its quite urgent... controller class coding /* * To change this template, choose Tools | Templates * and open the template in the editor. */ package project3; import java.net.URL; import java.util.ResourceBundle; import javafx.event.ActionEvent; import javafx.event.EventHandler; import javafx.fxml.FXML; import javafx.fxml.Initializable; import javafx.scene.control.Button; import javafx.scene.control.Label; import javafx.scene.control.TextField; import javafx.scene.layout.StackPane; import javafx.scene.shape.Circle; /** * FXML Controller class * * @xxx */ public class MyFxmlController implements Initializable { @FXML // fx:id="ball1" private Circle ball1; // Value injected by FXMLLoader @FXML // fx:id="ball2" private Circle ball2; // Value injected by FXMLLoader @FXML // fx:id="ball3" private Circle ball3; // Value injected by FXMLLoader @FXML // fx:id="ball4" private Circle ball4; // Value injected by FXMLLoader @FXML // fx:id="container" private Circle container; // Value injected by FXMLLoader @FXML // fx:id="myButton" private Button myButton; // Value injected by FXMLLoader @FXML // fx:id="myLabel1" private Label myLabel1; // Value injected by FXMLLoader @FXML // fx:id="myLabel2" private Label myLabel2; // Value injected by FXMLLoader @FXML // fx:id="pane" private StackPane pane; // Value injected by FXMLLoader @FXML // fx:id="txt" private TextField txt; // Value injected by FXMLLoader @Override // This method is called by the FXMLLoader when initialization is complete public void initialize(URL fxmlFileLocation, ResourceBundle resources) { assert ball1 != null : "fx:id=\"ball1\" was not injected: check your FXML file 'MyFxml.fxml'."; assert ball2 != null : "fx:id=\"ball2\" was not injected: check your FXML file 'MyFxml.fxml'."; assert ball3 != null : "fx:id=\"ball3\" was not injected: check your FXML file 'MyFxml.fxml'."; assert ball4 != null : "fx:id=\"ball4\" was not injected: check your FXML file 'MyFxml.fxml'."; assert container != null : "fx:id=\"container\" was not injected: check your FXML file 'MyFxml.fxml'."; assert myButton != null : "fx:id=\"myButton\" was not injected: check your FXML file 'MyFxml.fxml'."; assert myLabel1 != null : "fx:id=\"myLabel1\" was not injected: check your FXML file 'MyFxml.fxml'."; assert myLabel2 != null : "fx:id=\"myLabel2\" was not injected: check your FXML file 'MyFxml.fxml'."; assert pane != null : "fx:id=\"pane\" was not injected: check your FXML file 'MyFxml.fxml'."; assert txt != null : "fx:id=\"txt\" was not injected: check your FXML file 'MyFxml.fxml'."; // initialize your logic here: all @FXML variables will have been injected myButton.setOnAction(new EventHandler<ActionEvent>(){ @Override public void handle(ActionEvent event) { int count = 0; String guessed=txt.getText(); boolean result; result=MyCode.check(guessed); if(result) { ball1.setTranslateX(600); ball1.setTranslateY(250-container.getRadius()); //ball2.setTranslateX(600); // ball2.setTranslateY(250-container.getRadius()); } else System.out.println("wrong"); } }); } } word guessing logic public class MyCode { static String x="Netbeans"; static String y[]={"net","beans","neat","beat","bet"}; //static int counter; // public MyCode() { // counter++; //} static boolean check(String guessed) { int count=0; boolean result=false; //counter++; //System.out.println("turns"+counter); for(count=0;count<5;count++) { if(guessed.equals(y[count])) { result=true; break; } } if(result) System.out.println("Right"); else System.out.println("Wrong"); return result; } }

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  • Database Partitioning and Multiple Data Source Considerations

    - by Jeffrey McDaniel
    With the release of P6 Reporting Database 3.0 partitioning was added as a feature to help with performance and data management.  Careful investigation of requirements should be conducting prior to installation to help improve overall performance throughout the lifecycle of the data warehouse, preventing future maintenance that would result in data loss. Before installation try to determine how many data sources and partitions will be required along with the ranges.  In P6 Reporting Database 3.0 any adjustments outside of defaults must be made in the scripts and changes will require new ETL runs for each data source.  Considerations: 1. Standard Edition or Enterprise Edition of Oracle Database.   If you aren't using Oracle Enterprise Edition Database; the partitioning feature is not available. Multiple Data sources are only supported on Enterprise Edition of Oracle   Database. 2. Number of Data source Ids for partitioning during configuration.   This setting will specify how many partitions will be allocated for tables containing data source information.  This setting requires some evaluation prior to installation as       there are repercussions if you don't estimate correctly.   For example, if you configured the software for only 2 data sources and the partition setting was set to 2, however along came a 3rd data source.  The necessary steps to  accommodate this change are as follows: a) By default, 3 partitions are configured in the Reporting Database scripts. Edit the create_star_tables_part.sql script located in <installation directory>\star\scripts   and search for partition.  You’ll see P1, P2, P3.  Add additional partitions and sub-partitions for P4 and so on. These will appear in several areas.  (See P6 Reporting Database 3.0 Installation and Configuration guide for more information on this and how to adjust partition ranges). b) Run starETL -r.  This will recreate each table with the new partition key.  The effect of this step is that all tables data will be lost except for history related tables.   c) Run starETL for each of the 3 data sources (with the data source # (starETL.bat "-s2" -as defined in P6 Reporting Database 3.0 Installation and Configuration guide) The best strategy for this setting is to overestimate based on possible growth.  If during implementation it is deemed that there are atleast 2 data sources with possibility for growth, it is a better idea to set this setting to 4 or 5, allowing room for the future and preventing a ‘start over’ scenario. 3. The Number of Partitions and the Number of Months per Partitions are not specific to multi-data source.  These settings work in accordance to a sub partition of larger tables with regard to time related data.  These settings are dataset specific for optimization.  The number of months per partition is self explanatory, optimally the smaller the partition, the better query performance so if the dataset has an extremely large number of spread/history records, a lower number of months is optimal.  Working in accordance with this setting is the number of partitions, this will determine how many "buckets" will be created per the number of months setting.  For example, if you kept the default for # of partitions of 3, and select 2 months for each partitions you would end up with: -1st partition, 2 months -2nd partition, 2 months -3rd partition, all the remaining records Therefore with records to this setting, it is important to analyze your source db spread ranges and history settings when determining the proper number of months per partition and number of partitions to optimize performance.  Also be aware the DBA will need to monitor when these partition ranges will fill up and when additional partitions will need to be added.  If you get to the final range partition and there are no additional range partitions all data will be included into the last partition. 

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  • StreamInsight 2.1 Released

    - by Roman Schindlauer
    The wait is over—we are pleased to announce the release of StreamInsight 2.1. Since the release of version 1.2, we have heard your feedbacks and suggestions and based on that we have come up with a whole new set of features. Here are some of the highlights: A New Programming Model – A more clear and consistent object model, eliminating the need for complex input and output adapters (though they are still completely supported). This new model allows you to provision, name, and manage data sources and sinks in the StreamInsight server. Tight integration with Reactive Framework (Rx) – You can write reactive queries hosted inside StreamInsight as well as compose temporal queries on reactive objects. High Availability – Check-pointing over temporal streams and multiple processes with shared computation. Here is how simple coding can be with the 2.1 Programming Model: class Program {     static void Main(string[] args)     {         using (Server server = Server.Create("Default"))         {             // Create an app             Application app = server.CreateApplication("app");             // Define a simple observable which generates an integer every second             var source = app.DefineObservable(() =>                 Observable.Interval(TimeSpan.FromSeconds(1)));             // Define a sink.             var sink = app.DefineObserver(() =>                 Observer.Create<long>(x => Console.WriteLine(x)));             // Define a query to filter the events             var query = from e in source                         where e % 2 == 0                         select e;             // Bind the query to the sink and create a runnable process             using (IDisposable proc = query.Bind(sink).Run("MyProcess"))             {                 Console.WriteLine("Press a key to dispose the process...");                 Console.ReadKey();             }         }     } }   That’s how easily you can define a source, sink and compose a query and run it. Note that we did not replace the existing APIs, they co-exist with the new surface. Stay tuned, you will see a series of articles coming out over the next few weeks about the new features and how to use them. Come and grab it from our download center page and let us know what you think! You can find the updated MSDN documentation here, and we would appreciate if you could provide feedback to the docs as well—best via email to [email protected]. Moreover, we updated our samples to demonstrate the new programming surface. Regards, The StreamInsight Team

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  • StreamInsight 2.1 Released

    - by Roman Schindlauer
    The wait is over—we are pleased to announce the release of StreamInsight 2.1. Since the release of version 1.2, we have heard your feedbacks and suggestions and based on that we have come up with a whole new set of features. Here are some of the highlights: A New Programming Model – A more clear and consistent object model, eliminating the need for complex input and output adapters (though they are still completely supported). This new model allows you to provision, name, and manage data sources and sinks in the StreamInsight server. Tight integration with Reactive Framework (Rx) – You can write reactive queries hosted inside StreamInsight as well as compose temporal queries on reactive objects. High Availability – Check-pointing over temporal streams and multiple processes with shared computation. Here is how simple coding can be with the 2.1 Programming Model: class Program {     static void Main(string[] args)     {         using (Server server = Server.Create("Default"))         {             // Create an app             Application app = server.CreateApplication("app");             // Define a simple observable which generates an integer every second             var source = app.DefineObservable(() =>                 Observable.Interval(TimeSpan.FromSeconds(1)));             // Define a sink.             var sink = app.DefineObserver(() =>                 Observer.Create<long>(x => Console.WriteLine(x)));             // Define a query to filter the events             var query = from e in source                         where e % 2 == 0                         select e;             // Bind the query to the sink and create a runnable process             using (IDisposable proc = query.Bind(sink).Run("MyProcess"))             {                 Console.WriteLine("Press a key to dispose the process...");                 Console.ReadKey();             }         }     } }   That’s how easily you can define a source, sink and compose a query and run it. Note that we did not replace the existing APIs, they co-exist with the new surface. Stay tuned, you will see a series of articles coming out over the next few weeks about the new features and how to use them. Come and grab it from our download center page and let us know what you think! You can find the updated MSDN documentation here, and we would appreciate if you could provide feedback to the docs as well—best via email to [email protected]. Moreover, we updated our samples to demonstrate the new programming surface. Regards, The StreamInsight Team

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  • C++ strongly typed typedef

    - by Kian
    I've been trying to think of a way of declaring strongly typed typedefs, to catch a certain class of bugs in the compilation stage. It's often the case that I'll typedef an int into several types of ids, or a vector to position or velocity: typedef int EntityID; typedef int ModelID; typedef Vector3 Position; typedef Vector3 Velocity; This can make the intent of code more clear, but after a long night of coding one might make silly mistakes like comparing different kinds of ids, or adding a position to a velocity perhaps. EntityID eID; ModelID mID; if ( eID == mID ) // <- Compiler sees nothing wrong { /*bug*/ } Position p; Velocity v; Position newP = p + v; // bug, meant p + v*s but compiler sees nothing wrong Unfortunately, suggestions I've found for strongly typed typedefs include using boost, which at least for me isn't a possibility (I do have c++11 at least). So after a bit of thinking, I came upon this idea, and wanted to run it by someone. First, you declare the base type as a template. The template parameter isn't used for anything in the definition, however: template < typename T > class IDType { unsigned int m_id; public: IDType( unsigned int const& i_id ): m_id {i_id} {}; friend bool operator==<T>( IDType<T> const& i_lhs, IDType<T> const& i_rhs ); }; Friend functions actually need to be forward declared before the class definition, which requires a forward declaration of the template class. We then define all the members for the base type, just remembering that it's a template class. Finally, when we want to use it, we typedef it as: class EntityT; typedef IDType<EntityT> EntityID; class ModelT; typedef IDType<ModelT> ModelID; The types are now entirely separate. Functions that take an EntityID will throw a compiler error if you try to feed them a ModelID instead, for example. Aside from having to declare the base types as templates, with the issues that entails, it's also fairly compact. I was hoping anyone had comments or critiques about this idea? One issue that came to mind while writing this, in the case of positions and velocities for example, would be that I can't convert between types as freely as before. Where before multiplying a vector by a scalar would give another vector, so I could do: typedef float Time; typedef Vector3 Position; typedef Vector3 Velocity; Time t = 1.0f; Position p = { 0.0f }; Velocity v = { 1.0f, 0.0f, 0.0f }; Position newP = p + v*t; With my strongly typed typedef I'd have to tell the compiler that multypling a Velocity by a Time results in a Position. class TimeT; typedef Float<TimeT> Time; class PositionT; typedef Vector3<PositionT> Position; class VelocityT; typedef Vector3<VelocityT> Velocity; Time t = 1.0f; Position p = { 0.0f }; Velocity v = { 1.0f, 0.0f, 0.0f }; Position newP = p + v*t; // Compiler error To solve this, I think I'd have to specialize every conversion explicitly, which can be kind of a bother. On the other hand, this limitation can help prevent other kinds of errors (say, multiplying a Velocity by a Distance, perhaps, which wouldn't make sense in this domain). So I'm torn, and wondering if people have any opinions on my original issue, or my approach to solving it.

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  • Microsoft 2003 DNS sometimes cant query for some A pointers when their TTL expires

    - by Bq
    Warning Long question :) We have a win 2003 server with a DNS server, every now and then it cant provide us with some A pointers for a specific domain. I have a small script running which asks for SOA,NS and A records for the domain in question and sometimes when the TTL expires the DNS fails to get the A records again, a Clear Cache fixes the problem.. Have a look Here it worked when the TTL expired Thu Apr 29 15:24:20 METDST 2010 dig basefarm.net soa basefarm.net. 64908 IN SOA ns01.osl.basefarm.net. hostmaster.basefarm.net. 2010042613 86400 3600 2419200 600 ns01.osl.basefarm.net. 299 IN A 81.93.160.4 dig basefarm.net ns basefarm.net. 64908 IN NS ns01.sth.basefarm.net. basefarm.net. 64908 IN NS ns01.osl.basefarm.net. ns01.sth.basefarm.net. 299 IN A 80.76.149.76 ns01.osl.basefarm.net. 299 IN A 81.93.160.4 dig ns01.sth.basefarm.net a ns01.sth.basefarm.net. 299 IN A 80.76.149.76 The TTL expired for ns01.sth.basefarm.net and ns01.osl.basefarm.net but the DNS managed to get the new values (TTL 3600) Thu Apr 29 15:29:20 METDST 2010 dig basefarm.net soa basefarm.net. 64608 IN SOA ns01.osl.basefarm.net. hostmaster.basefarm.net. 2010042613 86400 3600 2419200 600 ns01.osl.basefarm.net. 3600 IN A 81.93.160.4 dig basefarm.net ns basefarm.net. 64608 IN NS ns01.sth.basefarm.net. basefarm.net. 64608 IN NS ns01.osl.basefarm.net. ns01.sth.basefarm.net. 3600 IN A 80.76.149.76 ns01.osl.basefarm.net. 3600 IN A 81.93.160.4 dig ns01.sth.basefarm.net a ns01.sth.basefarm.net. 3600 IN A 80.76.149.76 But then another time it fails, and we need to clear the dns cache for it to start working again... Thu Apr 29 17:24:23 METDST 2010 dig basefarm.net soa basefarm.net. 57705 IN SOA ns01.osl.basefarm.net. hostmaster.basefarm.net. 2010042613 86400 3600 2419200 600 ns01.osl.basefarm.net. 299 IN A 81.93.160.4 dig basefarm.net ns basefarm.net. 57705 IN NS ns01.sth.basefarm.net. basefarm.net. 57705 IN NS ns01.osl.basefarm.net. ns01.sth.basefarm.net. 299 IN A 80.76.149.76 ns01.osl.basefarm.net. 299 IN A 81.93.160.4 dig ns01.sth.basefarm.net a ns01.sth.basefarm.net. 299 IN A 80.76.149.76 The TTL expires but the DNS cant get the ip addresses for ns01.sth.basefarm.net and ns01.osl.basefarm.net Thu Apr 29 17:29:23 METDST 2010 dig basefarm.net soa basefarm.net. 57405 IN SOA ns01.osl.basefarm.net. hostmaster.basefarm.net. 2010042613 86400 3600 2419200 600 ns01.osl.basefarm.net. 3600 IN A 81.93.160.4 dig basefarm.net ns basefarm.net. 57405 IN NS ns01.sth.basefarm.net. basefarm.net. 57405 IN NS ns01.osl.basefarm.net. dig ns01.sth.basefarm.net a Lookup failed I'm really lost on this one and have tried asking Google but to no avail..

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  • Configuring Cisco 877W router from scratch for DHCP, WiFi, ADSL2+, NAT

    - by David M Williams
    Hi all, I apologise if this is a BIG question but I am quite lost with the Cisco IOS. I know what I want to achieve just not how to do it :( I have a Cisco 877W router with 4 FastEthernet interfaces, 1 ATM interface and 1 802.11 Radio. I want to set it up for a small network and am trying to construct a configuration below. I was using Google to try and flesh it out but I think I need help and guidance from actual experts! If it helps, output from show ver says Cisco IOS software, C870 software (C870-ADVSECURITYK9-M), version 12.4(4)T7, release software (fc1) ROM: System bootstrap, version 12.3(8r)YI4, release software Here's what I have so far, which hopefully outlines clearly enough what I am wanting to do. The bits in angle brackets are placeholders (eg the secret password). ! ! Set router hostname ! hostname Shazam ! ! Set usernames and passwords ! username david privilege 15 secret 0 <PASSWORD> enable secret <SECRETPASSWORD> ! ! Configure SSH and telnet access ! line vty 0 4 privilege level 15 login local transport input telnet ssh ! ! Local logging ! logging buffered 51200 warning ! ! Set date and time for NSW, Australia (GMT +10h) ! ! ! Set router IP address to 192.168.1.1 on FastEthernet0 port ! interface FastEthernet0 ip address 192.168.1.1 255.255.255.0 no shut ip nat inside ! ! Forward any unknown DNS requests to Google ! ip dns server ip name-server 8.8.8.8 ip name-server 8.8.4.4 ! ! Set up DHCP ! DHCP pool covers 192.168.1.100 - .199 ! Set gateway and DNS server to be the router, ie 192.168.1.1 ! service dhcp ip routing ip dhcp excluded-address 192.168.1.1 192.168.1.99 ip dhcp excluded-address 192.168.1.200 192.168.1.255 ip dhcp pool <DHCPPOOLNAME> network 192.168.1.0 255.255.255.0 default-router 192.168.1.1 dns-server 192.168.1.1 lease 7 ! ! DHCP reservations ! ! Assign IP address 192.168.1.105 to MAC address 00-21-5D-2F-58-04 ! ! Configure ADSL2 connection details ! interface atm dsl operating-mode adsl2+ ! ! Set up NAT rules ! ! Forward port 35394 to 192.168.1.105 ! ! Set up WiFi ! ! SSID visible, WPA2 security, Pre-shared key I'm hoping most of this is boiler-plate stuff to you guys. I'm keen to not just get a working script but to actually understand it also. Unfortunately, I'm finding the Cisco reference material online very complex. Thank you!

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  • Apple Airport Express, Extreme and Time Capsules, BT Home Hub, Wireless Extenders confusion

    - by Jamie Hartnoll
    I post quite frequently in Stack Overflow, but use Superuser less frequently. Mainly as I don't change hardware often and rarely have software issues! I live in a small stone cottage, and have an office in a separate building across a yard. I have a BT Homehub which is located in the cottage and a series of Ethernet cables running across the yard to the office. This is fine for my wired stuff. My main office computers are PCs running Windows 7 Ultimate, and one on Win7 Home, all working fine. I also have an old laptop on Win XP which works fine wirelessly in the house for those evenings in front of the TV catching up on a bit of work. I also have an iPhone and an iPad. Recently, I have been trying to get WiFi in the office so I can use Adobe Shadow (or whatever it now is!) to improve mobile web development efficiency using my iPhone and iPad, so I bought this: http://www.ebuyer.com/393462-zyxel-wre2205-500mbps-powerline-wireless-n300-range-extender-wre2205-gb0101f Thinking that would be lovely just plugged into the socket by the door in the office, extending the perimeter of the WiFi from my Homehub. I can't get it to work properly! If I plug a laptop into its ethernet port I can get it to connect to the Homehub and give me a kinda of wired, wireless extender. If, however, I plug the ethernet port into my home hub, it then seems to extend the network, but only my iOs devices work, and all my wired stuff stops working, and seems to create an infinite loop where windows connects to my homehob, and then rather to the internet, it then connects back to the extender thing. Anyway... in the meantime, I took a fatal trip to the Apple Store, where I purchased an Airport Express... solely for the purpose of hooking my iOs devices up as wireless music players in the house. I knew it had WiFi, but didn't want to use that part as an extender, I didn't think it would work on a Homehub anyway. It doesn't work on a Homehub! I now have a new wireless network in the house, which, when anything connects to it cannot connect to the Internet, so it works ONLY as a wireless music player. I then borrowed some Powerline Adaptors from someone and realised that this whole thing was getting totally out of control! It seems all the technology is out there but it's so complicated to get the right series of devices. To further add to the confusion, I wouldn't mind a network hard drive. I bought one that broke and lost everything, so now we're on to looking at the Apple Time Capsules. So my question is... IF... I buy an Apple Time Capsule, can I: Hook that up to my Homehub, leaving the homehub connected to the Internet so my Hub phones still work, then disable wireless on the homehub Link up my Airport Express to the Time Capsule PROPERLY so it will connect to the Internet Do the above with an Apple TV box should I buy one in future Use the Time Capsule as a network hard drive to store video and music that can be viewed/listened to via my iOS devices/Apple TV/Aiport Express anywhere even with my main PC off (this currently stores all this data) Hope that the IOS devices like the WiFi from the TimeCapsule better than the Homehub and work without extension, or buy another Airport Express to get WiFI in the office. Or... should I buy an Airport Extreme and use a USB hard drive for the network drive?

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  • Linux pptp client stops working after several hours

    - by Aron Rotteveel
    Here's the situation: Setup: 1 Windows Server 2008 machine acting as a Domain Controller and RRAS server 1 CentOS machine in a datacentre located elsewhere PPTP client running on CentOS machine, connected to the DC via When I connect to the DC, everything is working fine. I have set up a static IP for the dialup connection in my RRAS server so that the CentOS machine is automatically assigned the IP 192.168.1.240. Inside the VPN, it is not possible to access this machine on the local IP-address. Perfect. However, after several hours, it simply seems to stop working (IE: I cannot ping to or from this machine on the local network). The strange thing is, however: The DC shows the VPN client as still being connected The CentOS machine shows the network interface as being up There are no entries in my /var/log/messages that indicate a problem Output from ifconfig: ppp0 Link encap:Point-to-Point Protocol inet addr:192.168.1.240 P-t-P:192.168.1.160 Mask:255.255.255.255 UP POINTOPOINT RUNNING NOARP MULTICAST MTU:1396 Metric:1 RX packets:43 errors:0 dropped:0 overruns:0 frame:0 TX packets:58 errors:0 dropped:0 overruns:0 carrier:0 collisions:0 txqueuelen:3 RX bytes:4511 (4.4 KiB) TX bytes:15071 (14.7 KiB) Output from route -n: 192.168.1.160 0.0.0.0 255.255.255.255 UH 0 0 0 ppp0 192.168.1.0 0.0.0.0 255.255.255.0 U 0 0 0 ppp0 I have the following in my ip-up.local: route add -net 192.168.1.0 netmask 255.255.255.0 dev ppp0 The situation can be easily fixed by issueing a killall pppd and re-connecting. However, I obviously do not want to do this every X-hours or so. I have tried running pppd with both the debug as the kdebug flag but cannot find the cause of this problem. Currently, my ppp0 network interface seems to be running and the last log lines mentioning it are: Feb 19 14:10:40 graviton pppd[10934]: local IP address 192.168.1.240 Feb 19 14:10:40 graviton pppd[10934]: remote IP address 192.168.1.160 Feb 19 14:10:40 graviton pppd[10934]: Script /etc/ppp/ip-up started (pid 10952) Feb 19 14:10:40 graviton pppd[10934]: Script /etc/ppp/ip-up finished (pid 10952), status = 0x0 Feb 19 14:11:27 graviton pptp[10935]: anon log[decaps_gre:pptp_gre.c:414]: buffering packet 190 (expecting 189, lost or reordered) Feb 19 14:11:37 graviton pptp[10942]: anon log[logecho:pptp_ctrl.c:677]: Echo Request received. Feb 19 14:11:37 graviton pptp[10942]: anon log[ctrlp_rep:pptp_ctrl.c:251]: Sent control packet type is 6 'Echo-Reply' Feb 19 14:12:37 graviton pptp[10942]: anon log[logecho:pptp_ctrl.c:677]: Echo Request received. Feb 19 14:12:37 graviton pptp[10942]: anon log[ctrlp_rep:pptp_ctrl.c:251]: Sent control packet type is 6 'Echo-Reply' Feb 19 14:12:37 graviton pptp[10942]: anon log[logecho:pptp_ctrl.c:677]: Echo Reply received. Feb 19 14:13:37 graviton pptp[10942]: anon log[logecho:pptp_ctrl.c:677]: Echo Reply received. Feb 19 14:14:37 graviton pptp[10942]: anon log[logecho:pptp_ctrl.c:677]: Echo Reply received. Feb 19 14:15:37 graviton pptp[10942]: anon log[logecho:pptp_ctrl.c:677]: Echo Reply received. Feb 19 14:16:37 graviton pptp[10942]: anon log[logecho:pptp_ctrl.c:677]: Echo Reply received. Feb 19 14:19:37 graviton pptp[10942]: anon log[logecho:pptp_ctrl.c:677]: Echo Reply received. Feb 19 14:19:37 graviton pptp[10942]: anon log[logecho:pptp_ctrl.c:679]: no more Echo Reply/Request packets will be reported. I have enabled the persist option. The network interface is still running, but it is still impossible to send data through the VPN. Any help is appreciated.

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  • Setting up apache to view https pages

    - by zac
    I am trying to set up a site using vmware workstation, ubuntu 11.10, and apache2. The site works fine but now the https pages are not showing up. For example if I try to go to https://www.mysite.com/checkout I just see the message Not Found The requested URL /checkout/ was not found on this server. I dont really know what I am doing and have tried a lot of things to get the ssl certificates in there right. A few things I have in there, in my httpd.conf I just have : ServerName localhost In my ports.conf I have : NameVirtualHost *:80 Listen 80 <IfModule mod_ssl.c> # If you add NameVirtualHost *:443 here, you will also have to change # the VirtualHost statement in /etc/apache2/sites-available/default-ssl # to <VirtualHost *:443> # Server Name Indication for SSL named virtual hosts is currently not # supported by MSIE on Windows XP. Listen 443 http </IfModule> <IfModule mod_gnutls.c> Listen 443 http </IfModule> In the /etc/apache2/sites-available/default-ssl : <IfModule mod_ssl.c> <VirtualHost _default_:443> ServerAdmin webmaster@localhost DocumentRoot /var/www <Directory /> Options FollowSymLinks AllowOverride None </Directory> <Directory /var/www/> Options Indexes FollowSymLinks MultiViews AllowOverride None Order allow,deny allow from all </Directory> .... truncated in the sites-available/default I have : <VirtualHost *:80> ServerAdmin webmaster@localhost DocumentRoot /var/www #DocumentRoot /home/magento/site/ <Directory /> Options FollowSymLinks AllowOverride None </Directory> <Directory /var/www/> #<Directory /home/magento/site/> Options Indexes FollowSymLinks MultiViews AllowOverride None Order allow,deny allow from all </Directory> ScriptAlias /cgi-bin/ /usr/lib/cgi-bin/ <Directory "/usr/lib/cgi-bin"> AllowOverride None Options +ExecCGI -MultiViews +SymLinksIfOwnerMatch Order allow,deny Allow from all </Directory> ErrorLog ${APACHE_LOG_DIR}/error.log # Possible values include: debug, info, notice, warn, error, crit, # alert, emerg. LogLevel warn CustomLog ${APACHE_LOG_DIR}/access.log combined Alias /doc/ "/usr/share/doc/" <Directory "/usr/share/doc/"> Options Indexes MultiViews FollowSymLinks AllowOverride None Order deny,allow Deny from all Allow from 127.0.0.0/255.0.0.0 ::1/128 </Directory> </VirtualHost> <virtualhost *:443> SSLEngine on SSLCertificateFile /etc/apache2/ssl/server.crt SSLCertificateKeyFile /etc/apache2/ssl/server.key ServerAdmin webmaster@localhost <Directory /> Options FollowSymLinks AllowOverride None </Directory> <Directory /var/www/> #<Directory /home/magento/site/> Options Indexes FollowSymLinks MultiViews AllowOverride None Order allow,deny allow from all </Directory> </virtualhost> I also have in sites-availabe a file setup for my site url, www.mysite.com so in /etc/apache2/sites-available/mysite.com <VirtualHost *:80> ServerName mysite.com DocumentRoot /home/magento/mysite.com <Directory /> Options FollowSymLinks AllowOverride All </Directory> <Directory /home/magento/mysite.com/ > Options Indexes FollowSymLinks MultiViews AllowOverride All Order allow,deny allow from all </Directory> ErrorLog /home/magento/logs/apache.log # Possible values include: debug, info, notice, warn, error, crit, # alert, emerg. LogLevel warn </VirtualHost> <VirtualHost *:443> ServerName mysite.com DocumentRoot /home/magento/mysite.com <Directory /> Options FollowSymLinks AllowOverride All </Directory> <Directory /home/magento/mysite.com/ > Options Indexes FollowSymLinks MultiViews AllowOverride All Order allow,deny allow from all </Directory> ErrorLog /home/magento/logs/apache.log # Possible values include: debug, info, notice, warn, error, crit, # alert, emerg. LogLevel warn </VirtualHost> Thanks for any help getting this setup! As is probably obvious from this post I am pretty lost at this point.

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  • How do I achieve lossless JPEG joining without truncation of partial MCUs?

    - by Karan
    I am working on a project for which I need to join thousands of JPEG images losslessly (I'm not talking about the Lossless JPEG/JPEG 2000/JPEG-LS formats here). Aforementioned images have varying levels of chroma subsampling (1x1, 1x2, 2x1, 2x2), resulting in varying MCU sizes (8x8, 8x16, 16x8, 16x16 px). However, in any given set of images to be joined together, each image has identical characteristics. For now, let's assume I only have 2 images. Image #1 (I1) is 256x256px in size and #2 (I2) is 239x256px in size. 2x2 subsampling is used such that MCU size is 16x16px. I2 thus obviously has partial MCUs at the right edge, since its width is not evenly divisible by 16. (I've read that so-called 'partial' MCUs actually contain the data for a complete MCU, but the image dimensions instruct the renderer to only display the relevant pixels and ignore/hide the extra ones.) Looking around for tools that could help me accomplish this, I came across a modified version of JpegTran, that contains an experimental lossless crop 'n' drop (cut & paste) feature. All the other apps I encountered that support lossless JPEG editing seem to utilise IJG's (JpegTran) code, so this seemed to be the logical choice. Also, given the sheer number of images, I wanted something that could preferably be run from the command-line so that I could automate the process with a script. Unfortunately, while everything else worked fine, it seems JpegTran truncates the partial MCUs instead of retaining them. Thus in the example above, the final joined image contains all of I1, but only 224x256px of I2. Why 224? because 239 = 14x16+15, which means there are 14 full MCUs along the width, and 1 partial MCU (just 1px short of the complete 16px). The last 15px is what is getting blanked, leading to a 495x256px image with 15px of blank (grey) pixels at the right edge. See images below (shame that imgur re-compresses them): (left )+ (right) = As you can clearly see, the red portion (15px) of I2 has been truncated by JpegTran. If the MCUs were 8px in width, the lost portion would have been the right-most 7px of I2. Similarly, joining I3 (256x239px) *below * I1 would cause the loss of 7 or 15px, depending on the MCU height of course: (top) + (bottom) = If this is better suited to some other StackExchange (or even non-SE) site/forum where JPEG/image encoding experts hang out, do let me know. Can what I am attempting even be done, or is the so-called 'lossless' JPEG crop 'n' drop only valid for images with no partial MCUs? (Maybe that is why the feature is still in an "experimental state" more than a decade after being introduced...) Until I know for sure that it is impossible, I am not interested in suggestions for lossy joining. Avoiding any generational loss whatsoever is the sole reason why I'm breaking my head over this, else I'd have had this done and dusted ages ago. Also, I am not interested in suggestions related to switching image formats. I do not control the source of the images. If it can be done, how? Please keep in mind that any alternate apps suggested must ideally be capable of automation, given the requirements stated above. (But given how it's unlikely I'm even going to receive a useful answer given the constraints, I would be happy with any app suggestion just as long as it actually works. I can always look into an AutoIT/AHK script or something later to automate it.) I understand that an odd-sized final image might cause issues, so I am fully prepared to accept any solution, even if it results in blank (preferably black) padding pixels to the right/bottom. What I mean is, I don't care if I1 + I2 is 496x256px (1px padding) or even 512x256px (17px padding) in size, as long as the final image contains all the actual image data from both source images, and the entire process is lossless. Obviously the lesser the padding (if any), the better, but at this point any solution will do. A Windows-based solution would be perfect, but a Linux-based one would be entirely acceptable.

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  • Why do we get a sudden spike in response times?

    - by Christian Hagelid
    We have an API that is implemented using ServiceStack which is hosted in IIS. While performing load testing of the API we discovered that the response times are good but that they deteriorate rapidly as soon as we hit about 3,500 concurrent users per server. We have two servers and when hitting them with 7,000 users the average response times sit below 500ms for all endpoints. The boxes are behind a load balancer so we get 3,500 concurrents per server. However as soon as we increase the number of total concurrent users we see a significant increase in response times. Increasing the concurrent users to 5,000 per server gives us an average response time per endpoint of around 7 seconds. The memory and CPU on the servers are quite low, both while the response times are good and when after they deteriorate. At peak with 10,000 concurrent users the CPU averages just below 50% and the RAM sits around 3-4 GB out of 16. This leaves us thinking that we are hitting some kind of limit somewhere. The below screenshot shows some key counters in perfmon during a load test with a total of 10,000 concurrent users. The highlighted counter is requests/second. To the right of the screenshot you can see the requests per second graph becoming really erratic. This is the main indicator for slow response times. As soon as we see this pattern we notice slow response times in the load test. How do we go about troubleshooting this performance issue? We are trying to identify if this is a coding issue or a configuration issue. Are there any settings in web.config or IIS that could explain this behaviour? The application pool is running .NET v4.0 and the IIS version is 7.5. The only change we have made from the default settings is to update the application pool Queue Length value from 1,000 to 5,000. We have also added the following config settings to the Aspnet.config file: <system.web> <applicationPool maxConcurrentRequestsPerCPU="5000" maxConcurrentThreadsPerCPU="0" requestQueueLimit="5000" /> </system.web> More details: The purpose of the API is to combine data from various external sources and return as JSON. It is currently using an InMemory cache implementation to cache individual external calls at the data layer. The first request to a resource will fetch all data required and any subsequent requests for the same resource will get results from the cache. We have a 'cache runner' that is implemented as a background process that updates the information in the cache at certain set intervals. We have added locking around the code that fetches data from the external resources. We have also implemented the services to fetch the data from the external sources in an asynchronous fashion so that the endpoint should only be as slow as the slowest external call (unless we have data in the cache of course). This is done using the System.Threading.Tasks.Task class. Could we be hitting a limitation in terms of number of threads available to the process?

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  • SQL Server Issue: Could not allocate space for object ... primary filegroup is full

    - by Luke
    Trying to figure out a problem at an office that has SQL Server 2005 installed on Windows SBS Server 2008. Here's the setup: It's an office, and the person who set this all up is nowhere to be found. I'm the best hope they have... One of the programs they use on a workstation gives them an error of "Could not allocate space for object 'Billing' in database "MyDatabase" because primary filegroup is full" when trying to save an entry in their software. I searched around for hours, looking for possible solutions. One was to check for available disk space, and another was to defrag. I checked the hard drives on the server, and there is plenty of space free. I also defragged, which may have helped the problem somewhat. It's hard to say, because it seems like with the nature of the error, if you try over and over you might get it to actually save. My next step was to try to see if autogrowth was enabled on the database. This would seem to be a likely / possible solution, but I can't access the database! If I run the SQL Management Studio, I can log in as my Windows user and view the list of databases. However, if I try to do anything (actually view the database, view the properties, add or edit users), I get errors that I don't have permission. For what it's worth, I also tried runing Management Studio as Administrator, in case that would help. No difference, though. Now, what I'm guessing is going on -- from my limited knowledge of SQL and from reading online -- is that though I'm logged in as a Windows administrator, that account does NOT have SQL access. I do see a list of SQL users, including SA, but I again don't have permission to add one or to change the password on an existing one. And nobody at the office has any idea what the SQL passwords could be. So... here's my thinking thus far: 1 - The "Could not allocate" error likely points to a database that needs to be allowed to autogrow. Especially since I verified there is plenty of free space and the HD has been defragmented. 2 - Enabling autogrow would be very easy to do if I had the proper access within SQL Management Stuido. That leads me to this link: http://blogs.technet.com/b/sqlman/archive/2011/06/14/tips-amp-tricks-you-have-lost-access-to-sql-server-now-what.aspx It sounds like it's a step-by-step guide for giving me the access I need to SQL. I'm guessing that if I followed this guide, I would be able to then log in to the SQL server via Management Studio with the proper permissions, and would be able to enable autogrow (or simply view the status of the existing database), and hopefully solve the "Could not allocate space" problem! So I guess I have a few questions: 1 - Would you guys agree with my "diagnosis"? Think I'm barking up the right tree? 2 - Is there any risk at all in hurting / disabling / wrecking the current SQL database or setup with me going through the guide to regain SQL access? I understand that per the guide, I would have to temporarily shut down SQL, so obviously it wouldn't be accessible during that time. But it wouldn't be worth the risk if there's a chance I could mess anything up... Like I said, the workstations ARE currently accessing the database somehow, but nobody knows with what login info or anything. Basically, it's set up, it works (usually), but if they had to reload the software, nobody would know how. Any feedback would be appreciated!! The problem is such that it's not an emergency for them, but an annoyance. If I could fix it, it would be wonderful. But if not, I think they'll manage, especially as they are going to eventually stop using this software. Thank you so much for your time! Luke

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  • SATA drive problems with two SIL RAID cards

    - by Jon Topper
    I've just put a second SiI 3114 SATARaid card in my home server so that I could add another pair of SATA drives and increase my storage space. Annoyingly, it doesn't seem to work: [ 32.816030] ata5: lost interrupt (Status 0x0) [ 32.816072] ata5.00: exception Emask 0x0 SAct 0x0 SErr 0x0 action 0x6 frozen [ 32.816091] ata5.00: cmd c8/00:08:00:00:00/00:00:00:00:00/e0 tag 0 dma 4096 in [ 32.816094] res 40/00:00:00:00:00/00:00:00:00:00/00 Emask 0x4 (timeout) [ 32.816101] ata5.00: status: { DRDY } [ 32.816117] ata5: hard resetting link [ 33.136082] ata5: SATA link down (SStatus 0 SControl 0) [ 36.060940] irq 18: nobody cared (try booting with the "irqpoll" option) [ 36.060949] Pid: 0, comm: swapper Not tainted 2.6.31-20-generic #58-Ubuntu [ 36.060954] Call Trace: [ 36.060977] [] ? printk+0x18/0x1c [ 36.060997] [] __report_bad_irq+0x27/0x90 [ 36.061005] [] note_interrupt+0x150/0x190 [ 36.061011] [] handle_fasteoi_irq+0xac/0xd0 [ 36.061023] [] handle_irq+0x18/0x30 [ 36.061029] [] do_IRQ+0x47/0xc0 [ 36.061042] [] ? irq_exit+0x50/0x70 [ 36.061058] [] ? smp_apic_timer_interrupt+0x57/0x90 [ 36.061065] [] common_interrupt+0x30/0x40 [ 36.061075] [] ? native_safe_halt+0x5/0x10 [ 36.061082] [] default_idle+0x46/0xd0 [ 36.061088] [] cpu_idle+0x8c/0xd0 [ 36.061103] [] rest_init+0x55/0x60 [ 36.061111] [] start_kernel+0x2e6/0x2ec [ 36.061117] [] ? unknown_bootoption+0x0/0x19e [ 36.061133] [] i386_start_kernel+0x7c/0x83 [ 36.061137] handlers: [ 36.061139] [] (sil_interrupt+0x0/0xb0) [ 36.061151] Disabling IRQ #18 [ 38.136014] ata5: hard resetting link [ 38.456022] ata5: SATA link down (SStatus 0 SControl 0) [ 43.456013] ata5: hard resetting link [ 43.776022] ata5: SATA link down (SStatus 0 SControl 0) [ 43.776035] ata5.00: disabled [ 43.776055] ata5.00: device reported invalid CHS sector 0 [ 43.776074] sd 4:0:0:0: [sde] Result: hostbyte=DID_OK driverbyte=DRIVER_SENSE [ 43.776082] sd 4:0:0:0: [sde] Sense Key : Aborted Command [current] [descriptor] [ 43.776092] Descriptor sense data with sense descriptors (in hex): [ 43.776097] 72 0b 00 00 00 00 00 0c 00 0a 80 00 00 00 00 00 [ 43.776112] 00 00 00 00 [ 43.776118] sd 4:0:0:0: [sde] Add. Sense: No additional sense information [ 43.776127] end_request: I/O error, dev sde, sector 0 [ 43.776136] Buffer I/O error on device sde, logical block 0 [ 43.776170] ata5: EH complete [ 43.776187] ata5.00: detaching (SCSI 4:0:0:0) root@core:~# cat /proc/interrupts CPU0 0: 47 IO-APIC-edge timer 1: 8 IO-APIC-edge i8042 6: 3 IO-APIC-edge floppy 7: 0 IO-APIC-edge parport0 8: 0 IO-APIC-edge rtc0 9: 0 IO-APIC-fasteoi acpi 14: 53069 IO-APIC-edge pata_sis 15: 53004 IO-APIC-edge pata_sis 17: 112265 IO-APIC-fasteoi sata_sil 18: 200002 IO-APIC-fasteoi sata_sil, SiS SI7012 19: 111140 IO-APIC-fasteoi eth0 20: 0 IO-APIC-fasteoi ohci_hcd:usb2 21: 0 IO-APIC-fasteoi ohci_hcd:usb3 23: 0 IO-APIC-fasteoi ehci_hcd:usb1 NMI: 0 Non-maskable interrupts LOC: 6650492 Local timer interrupts SPU: 0 Spurious interrupts CNT: 0 Performance counter interrupts PND: 0 Performance pending work RES: 0 Rescheduling interrupts CAL: 0 Function call interrupts TLB: 0 TLB shootdowns TRM: 0 Thermal event interrupts THR: 0 Threshold APIC interrupts MCE: 0 Machine check exceptions MCP: 160 Machine check polls ERR: 0 MIS: 0 root@core:~# lspci | grep Raid 00:09.0 RAID bus controller: Silicon Image, Inc. SiI 3114 [SATALink/SATARaid] Serial ATA Controller (rev 02) 00:0a.0 RAID bus controller: Silicon Image, Inc. SiI 3114 [SATALink/SATARaid] Serial ATA Controller (rev 02) root@core:~# lsb_release -a No LSB modules are available. Distributor ID: Ubuntu Description: Ubuntu 9.10 Release: 9.10 Codename: karmic root@core:~# uname -a Linux core.topper.me.uk 2.6.31-20-generic #58-Ubuntu SMP Fri Mar 12 05:23:09 UTC 2010 i686 GNU/Linux I've tried a combination of different kernel options (irqpoll, noapic, noacpi, pci=noapic) all to no avail. Does anyone have any bright ideas about how I can go about making this work? Swapping PCI cards around isn't an option as there are only two slots in this motherboard (an ASRock K7S41GX). The BIOS doesn't look to have too much in the way of configuration options regarding IRQ usage. Plan B is to ditch this server completely and buy a new QNAP for these drives to go in, but I was hoping to avoid doing this right now.

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  • How to get rid of a stubborn 'removed' device in mdadm

    - by T.J. Crowder
    One of my server's drives failed and so I removed the failed drive from all three relevant arrays, had the drive swapped out, and then added the new drive to the arrays. Two of the arrays worked perfectly. The third added the drive back as a spare, and there's an odd "removed" entry in the mdadm details. I tried both mdadm /dev/md2 --remove failed and mdadm /dev/md2 --remove detached as suggested here and here, neither of which complained, but neither of which had any effect, either. Does anyone know how I can get rid of that entry and get the drive added back properly? (Ideally without resyncing a third time, I've already had to do it twice and it takes hours. But if that's what it takes, that's what it takes.) The new drive is /dev/sda, the relevant partition is /dev/sda3. Here's the detail on the array: # mdadm --detail /dev/md2 /dev/md2: Version : 0.90 Creation Time : Wed Oct 26 12:27:49 2011 Raid Level : raid1 Array Size : 729952192 (696.14 GiB 747.47 GB) Used Dev Size : 729952192 (696.14 GiB 747.47 GB) Raid Devices : 2 Total Devices : 2 Preferred Minor : 2 Persistence : Superblock is persistent Update Time : Tue Nov 12 17:48:53 2013 State : clean, degraded Active Devices : 1 Working Devices : 2 Failed Devices : 0 Spare Devices : 1 UUID : 2fdbf68c:d572d905:776c2c25:004bd7b2 (local to host blah) Events : 0.34665 Number Major Minor RaidDevice State 0 0 0 0 removed 1 8 19 1 active sync /dev/sdb3 2 8 3 - spare /dev/sda3 If it's relevant, it's a 64-bit server. It normally runs Ubuntu, but right now I'm in the data centre's "rescue" OS, which is Debian 7 (wheezy). The "removed" entry was there the last time I was in Ubuntu (it won't, currently, boot from the disk), so I don't think that's not some Ubuntu/Debian conflict (and they are, of course, closely related). Update: Having done extensive tests with test devices on a local machine, I'm just plain getting anomalous behavior from mdadm with this array. For instance, with /dev/sda3 removed from the array again, I did this: mdadm /dev/md2 --grow --force --raid-devices=1 And that got rid of the "removed" device, leaving me just with /dev/sdb3. Then I nuked /dev/sda3 (wrote a file system to it, so it didn't have the raid fs anymore), then: mdadm /dev/md2 --grow --raid-devices=2 ...which gave me an array with /dev/sdb3 in slot 0 and "removed" in slot 1 as you'd expect. Then mdadm /dev/md2 --add /dev/sda3 ...added it — as a spare again. (Another 3.5 hours down the drain.) So with the rebuilt spare in the array, given that mdadm's man page says RAID-DEVICES CHANGES ... When the number of devices is increased, any hot spares that are present will be activated immediately. ...I grew the array to three devices, to try to activate the "spare": mdadm /dev/md2 --grow --raid-devices=3 What did I get? Two "removed" devices, and the spare. And yet when I do this with a test array, I don't get this behavior. So I nuked /dev/sda3 again, used it to create a brand-new array, and am copying the data from the old array to the new one: rsync -r -t -v --exclude 'lost+found' --progress /mnt/oldarray/* /mnt/newarray This will, of course, take hours. Hopefully when I'm done, I can stop the old array entirely, nuke /dev/sdb3, and add it to the new array. Hopefully, it won't get added as a spare!

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  • Unmountable boot volume blue screen, what should I do?

    - by Josh
    I was trying to install an update from NVIDIA for my GTX 560, but while it was installing, my computer shut off. After a few minutes, I turned it back on. It got to the Windows boot screen and then had a blue screen error and if left on it would just keep doing that. A few details about my PC: I haven't added any new hardware or software, I'm running Windows XP Professional 32 bit and Windows XP Professional 64 bit on the same hard drive for about 2 years now. I have 2 other hard drives also, but I don't have one large enough to hold everything from my main hard drive, so formatting isn't an option. Now, as for what I've done so far: I've scanned the RAM with "memtest - 86 v3.4" and it said that it was good. I scanned the hard drive in question with chkdsk /r and it gets to 50% and tells me something along the lines of "the drive has one or more unrepairable problems". I also tried to use chkdsk on the drive I installed the new copy of Windows XP on and it got to 75% then jumped back down to 50% and stayed there (I had to reboot the pc). So, after that, I turned off auto reboot and got to read the blue screen error code and I looked it up only to find that nobody seems to have this problem, just problems close to it. The error code is 0x000000ed and I've seen a lot of these online but none that matched the detailed part of the code UNMOUNTABLE_BOOT_VOLUME 0x000000ed (0xfffffadf513c19a0, 0xffffffffc0000006, 0, 0) So, I have installed another copy of Windows XP Professional 32 bit on one of my other hard drives in hopes of accessing the data on the drive in question and when it booted it asked if I wanted chkdsk to scan the drive in question and this is what it found: file record segments 12740, 12741, 12742 and 12743 were reported unreadable. Then it says "recovering lost files" but it sits there for a few seconds and then just boots to Windows. I can't access the drive in question from Windows as far as I can tell, it just says "drive not accessible" and when I go to properties it says that the drive has 100% free space. So, after that failed I didn't give up, I looked for another way to access the drive in question. I used a Ubuntu bootable disk and was able to access the drive in question without any problems. However, I can't access the registry editor because it's a .exe file and that won't load from Ubuntu. I made a copy of the "Windows" folder and put it on one of my other drives and that's where I'm stuck at now. I'm sure my drive works fine, I know chkdsk can't fix the problem with it and I know what caused the problem in the first place for the most part, but I don't know what to do about it. I have a laptop that I can use to download and burn disks if needed and I also have the other copy of Windows XP Professional 32 bit that I can use that's installed on the computer in question (so I know it's not a hardware issue). I'm pretty sure it's a driver issue or the update was editing the registry when it shut off and left me when a broken registry. I've tried accessing C:\Windows\System32\CONFIG only to find that the Windows XP disk repair option can't even access the files on the drive in question. It seems I'll need to be able to do everything from Ubuntu unless there is something I haven't tried with the Windows XP disk. I didn't install the update on Windows XP 64 bit but yet it also has the same blue screen error (that's where the error code above came from but I haven't checked to see if they are the same). They both stopped working at the same time, so I assume it's one problem causing both to not work.

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  • Securing smtp with login

    - by Paul Peelen
    I have a ispconfig server, and it seems that someone is using it to send spam. I got about 130 "Mail Delivery System" email about declined send email. This spammer uses my email address as sent from adress, so I get all these email adresses to my mail. I am using Postfix and Courier. I installed my server according to this guide: http://www.howtoforge.com/perfect-server-debian-lenny-ispconfig3-p3 I did this a few months ago. My question: Can I secure my server to require login to be able to send email, and if so... how? Thanks! EDIT Some data from mail.log, these kind of error show up constantly: Jun 15 17:58:16 bolt postfix/qmgr[10712]: CC7DA1242AE: from=<paul@*****.se>, size=3782, nrcpt=1 (queue active) Jun 15 17:58:16 bolt postfix/smtp[11337]: CC7DA1242AE: to=<[email protected]>, relay=none, delay=4641, delays=4640/0.01/0.32/0, dsn=4.4.3, status=deferred (Host or domain name not found. Name service error for name=cmlisboa.pt type=MX: Host not found, try again) Jun 15 17:58:19 bolt postfix/smtpd[10836]: connect from static-200-105-220-154.acelerate.net[200.105.220.154] Jun 15 17:58:20 bolt postfix/smtpd[10836]: NOQUEUE: reject: RCPT from static-200-105-220-154.acelerate.net[200.105.220.154]: 550 5.1.1 <advertising@*****.com>: Recipient address rejected: User unknown in virtual mailbox table; from=<[email protected]> to=<advertising@*****.com> proto=ESMTP helo=<static-200-105-220-154.acelerate.net> Jun 15 17:58:20 bolt postfix/smtpd[10836]: lost connection after DATA (0 bytes) from static-200-105-220-154.acelerate.net[200.105.220.154] Jun 15 17:58:20 bolt postfix/smtpd[10836]: disconnect from static-200-105-220-154.acelerate.net[200.105.220.154] Jun 15 17:58:29 bolt postfix/smtpd[10834]: connect from unknown[62.176.172.226] Jun 15 17:58:32 bolt postfix/smtpd[10834]: 386791241F9: client=unknown[62.176.172.226] Jun 15 17:58:34 bolt postfix/cleanup[10975]: 386791241F9: message-id=<[email protected]> Jun 15 17:58:34 bolt postfix/qmgr[10712]: 386791241F9: from=<[email protected]>, size=867, nrcpt=1 (queue active) Jun 15 17:58:35 bolt postfix/smtpd[10834]: disconnect from unknown[62.176.172.226] Jun 15 17:58:35 bolt amavis[11084]: (11084-17) Blocked SPAM, [62.176.172.226] [62.176.172.226] <[email protected]> -> <*****@*****>, Message-ID: <[email protected]>, mail_id: XczovKoMBYNr, Hits: 18.471, size: 867, 833 ms Jun 15 17:58:35 bolt postfix/smtp[10732]: 386791241F9: to=<*****@*****>, relay=127.0.0.1[127.0.0.1]:10024, delay=3.5, delays=2.7/0/0/0.83, dsn=2.7.0, status=sent (250 2.7.0 Ok, discarded, id=11084-17 - SPAM) Jun 15 17:58:35 bolt postfix/qmgr[10712]: 386791241F9: removed Jun 15 17:58:43 bolt postfix/smtpd[10836]: warning: 178.121.154.194: address not listed for hostname mm-194-154-121-178.dynamic.pppoe.mgts.by Jun 15 17:58:43 bolt postfix/smtpd[10836]: connect from unknown[178.121.154.194] Jun 15 17:58:45 bolt postfix/smtpd[10727]: connect from unknown[180.134.223.86] EDIT #2 Got some more info from the logs, this is a send request: mail.info.1:Jun 15 16:41:57 bolt amavis[5399]: (05399-06) Passed CLEAN, [110.139.48.64] [110.139.48.64] <paul@*****.se> -> <[email protected]>, Message-ID: <CHILKAT-MID-7c54ebcf-5501-de9b-f0b1-4f0234290d8d@HP-IRISH>, mail_id: 35l56Ramx6Nc, Hits: -2.941, size: 3329, queued_as: 2485770086, 136 ms mail.info.1:Jun 15 16:41:57 bolt postfix/smtp[4743]: 375C570082: to=<[email protected]>, relay=127.0.0.1[127.0.0.1]:10024, delay=4.8, delays=4.7/0/0/0.14, dsn=2.0.0, status=sent (250 2.0.0 Ok, id=05399-06, from MTA([127.0.0.1]:10025): 250 2.0.0 Ok: queued as 2485770086) Which apparently got thrue. Any ideas how to restrict this?

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  • Courier Maildrop error user unknown. Command output: Invalid user specified

    - by cad
    Hello I have a problem with maildrop. I have read dozens of webs/howto/emails but couldnt solve it. My objective is moving automatically spam messages to a spam folder. My email server is working perfectly. It marks spam in subject and headers using spamassasin. My box has: Ubuntu 9.04 Web: Apache2 + Php5 + MySQL MTA: Postfix 2.5.5 + SpamAssasin + virtual users using mysql IMAP: Courier 0.61.2 + Courier AuthLib WebMail: SquirrelMail I have read that I could use Squirrelmail directly (not a good idea), procmail or maildrop. As I already have maildrop in the box (from courier) I have configured the server to use maildrop (added an entry in transport table for a virtual domain). I found this error in email: This is the mail system at host foo.net I'm sorry to have to inform you that your message could not be delivered to one or more recipients. It's attached below. For further assistance, please send mail to postmaster. If you do so, please include this problem report. You can delete your own text from the attached returned message. The mail system <[email protected]>: user unknown. Command output: Invalid user specified. Final-Recipient: rfc822; [email protected] Action: failed Status: 5.1.1 Diagnostic-Code: x-unix; Invalid user specified. ---------- Forwarded message ---------- From: test <[email protected]> To: [email protected] Date: Sat, 1 May 2010 19:49:57 +0100 Subject: fail fail An this in the logs May 1 18:50:18 foo.net postfix/smtpd[14638]: connect from mail-bw0-f212.google.com[209.85.218.212] May 1 18:50:19 foo.net postfix/smtpd[14638]: 8A9E9DC23F: client=mail-bw0-f212.google.com[209.85.218.212] May 1 18:50:19 foo.net postfix/cleanup[14643]: 8A9E9DC23F: message-id=<[email protected]> May 1 18:50:19 foo.net postfix/qmgr[14628]: 8A9E9DC23F: from=<[email protected]>, size=1858, nrcpt=1 (queue active) May 1 18:50:23 foo.net postfix/pickup[14627]: 1D4B4DC2AA: uid=5002 from=<[email protected]> May 1 18:50:23 foo.net postfix/cleanup[14643]: 1D4B4DC2AA: message-id=<[email protected]> May 1 18:50:23 foo.net postfix/pipe[14644]: 8A9E9DC23F: to=<[email protected]>, relay=spamassassin, delay=3.8, delays=0.55/0.02/0/3.2, dsn=2.0.0, status=sent (delivered via spamassassin service) May 1 18:50:23 foo.net postfix/qmgr[14628]: 8A9E9DC23F: removed May 1 18:50:23 foo.net postfix/qmgr[14628]: 1D4B4DC2AA: from=<[email protected]>, size=2173, nrcpt=1 (queue active) **May 1 18:50:23 foo.netpostfix/pipe[14648]: 1D4B4DC2AA: to=<[email protected]>, relay=maildrop, delay=0.22, delays=0.06/0.01/0/0.15, dsn=5.1.1, status=bounced (user unknown. Command output: Invalid user specified. )** May 1 18:50:23 foo.net postfix/cleanup[14643]: 4C2BFDC240: message-id=<[email protected]> May 1 18:50:23 foo.net postfix/qmgr[14628]: 4C2BFDC240: from=<>, size=3822, nrcpt=1 (queue active) May 1 18:50:23 foo.net postfix/bounce[14651]: 1D4B4DC2AA: sender non-delivery notification: 4C2BFDC240 May 1 18:50:23 foo.net postfix/qmgr[14628]: 1D4B4DC2AA: removed May 1 18:50:24 foo.net postfix/smtp[14653]: 4C2BFDC240: to=<[email protected]>, relay=gmail-smtp-in.l.google.com[209.85.211.97]:25, delay=0.91, delays=0.02/0.03/0.12/0.74, dsn=2.0.0, status=sent (250 2.0.0 OK 1272739824 37si5422420ywh.59) May 1 18:50:24 foo.net postfix/qmgr[14628]: 4C2BFDC240: removed My config files: http://lar3d.net/main.cf (/etc/postfix) http://lar3d.net/master.c (/etc/postfix) http://lar3d.net/local.cf (/etc/spamassasin) http://lar3d.net/maildroprc (maildroprc) If I change master.cf line (as suggested here) maildrop unix - n n - - pipe flags=DRhu user=vmail argv=/usr/lib/courier/bin/maildrop -d ${recipient} with maildrop unix - n n - - pipe flags=DRhu user=vmail argv=/usr/lib/courier/bin/maildrop -d vmail ${recipient} I get the email in /home/vmail/MailDir instead of the correct dir (/home/vmail/foo.net/info/.SPAM ) After reading a lot I have some guess but not sure. - Maybe I have to install userdb? - Maybe is something related with mysql, but everything is working ok - If I try with procmail I will face same problem... - What are flags DRhu for? Couldnt find doc about them - In some places I found maildrop line with more parameters flags=DRhu user=vmail argv=/usr/lib/courier/bin/maildrop -d $ ${recipient} ${extension} ${recipient} ${user} ${nexthop} ${sender} I am really lost. Dont know how to continue. If you have any idea or need another config file please let me know. Thanks!!!

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