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  • Can I pass an array as arguments to a method with variable arguments in Java?

    - by user352382
    I'd like to be able to create a function like: class A { private String extraVar; public String myFormat(String format, Object ... args){ return String.format(format, extraVar, args); } } The problem here is that args is treated as Object[] in the method myFormat, and thus is a single argument to String.format, while I'd like every single Object in args to be passed as a new argument. Since String.format is also a method with variable arguments, this should be possible. If this is not possible, is there a method like String.format(String format, Object[] args)? In that case I could prepend extraVar to args using a new array and pass it to that method.

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  • private virtual function in derived class

    - by user1706047
    class base { public: virtual void doSomething() = 0; }; class derived : public base { **private:** virtual void doSomething(){cout<<"Derived fn"<<endl;} }; now if i do the following: base *b=new child; b->doSomething(); //it calls the derived class fn even if that is private. Question: 1.its able to call the derived class fn even if that is private.How is it possible? Now if i change the inheritance access specifier from public to protected/private then i get compilation error as "'type cast' : conversion from 'Derived *' to 'base *' exists, but is inaccessible" Notes: I am aware on the concepts of the inheritance access specifiers.So in second case as its derived private/protected, its inaccessible. But here it confuses me for the first question. Any input will be highly appreciated

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  • CSS: How to set remaining width as %, but having knowledge of only pixels

    - by Mega Matt
    Hi all, I've seen this question asked in a couple other contexts on SO, but I thought it would be worth asking again for my particular case. I'm trying to create some re-usable CSS classes for more consistency and less clutter on my site, and I'm stuck on trying to standardize one thing I use frequently. I have a container div that I don't want to set the height for (because it will vary depending on where on the site it is), and inside it is a header div, and then an unordered list of items, all with CSS applied to them. It looks a lot like this: I want the unordered list to take up the remaining room in the container div, knowing that the header div is 18px tall. I just don't know how to specify the list's height as "the result of 100% minus 18px". Does anyone have any advice in this situation? Thanks very much.

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  • Why is it called NoSQL?

    - by beef jerky
    I've recently worked with MongoDB and learned about its schemaless design. However, I'm confused with the term NoSQL? Why is it called that? Doesn't it use SQL or SQL-like queries? I've also read from an article that the main difference lies in how data is stored. In the case of MongoDB, it's stored like JSON documents. Is this true? Also, I'm confused why I always see 'NoSQL vs relational databases'. Aren't NoSQL databases relational? I believe documents in MongoDB are still related/linked through some keys (please correct me if I'm wrong). So why is it labeled as non-relational? Thanks in advance!

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  • JQuery mobile handling links (return false)

    - by shinax
    I'm planning on using JQM to make a simple mobile web app, and I'm having problems getting this simple functionality to work: when I click on some links (not all of them), I want to be able to first process some data, and then depending on that outcome, sometimes continue the link action (I like the whole transitions and ajax things), and sometimes don't. The important part is that I want to preserve the normal JQMobile transitions and stuff for the links, just sometimes prevent them (for example for validation and things like that). I've tried with: return false, preventDefault (and in combination with stopPropagation), and data-ajax="false", and none hav worked, they all redirect. Could somebody tell me the correct way to to this in JQMobile? Just in case it's important: I'm using anchor links, using this to test. Thank you in advance, Jennifer.

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  • Module.new with class_eval

    - by dorelal
    This is a large commit. But I want you to concentrate on this change block. http://github.com/rails/rails/commit/d916c62cfc7c59ab6411407a05b946d3dd7535e9#L2L1304 Even without understanding the full context of the code I am not able to think of a scenario where I would use include Modue.new { class_eval <<-RUBY def foo puts 'foo' end RUBY } Then end result is that in the root context (self just before include Moduel.new) a method call foo has been added. If I take out the Module.new code and if I only leave class_eval in that case also I will have a method called foo in self. What am I missing.

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  • Rotating a movieclip slowly (so it looks like it was animated) [on hold]

    - by user2537021
    Im trying to rotate an object X degrees, nut i want that when the rotation happends it isent all suden, i want that you can see the object spining all the way to the X degree. (im try ing to me more clear) (I'm pretty new at this so please be patient with me) switch (enemywalk) { case 1: colx = enemy.y; enemy.y -= 115; if ( enemy.hitTestObject (pared1) || enemy.hitTestObject (pared2) ) { enemy.y = colx; } enemy.rotation = enemyrotate; //random generated number break; }

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  • Why can't I decrease the line-height of this text?

    - by daGUY
    http://jsfiddle.net/mJxn4/ This is very odd: I have a few lines of text wrapped in an <em> tag. No matter what I do, lowering the value for line-height below 17px has no effect. I can bump the line-height up to greater than 17px and it'll apply, but I can't get it lower than 17px. The CSS in question is: #others .item em { font-size: 13px; line-height: 17px; } Try adjusting the line height both higher and lower and run the updated fiddle after each change, and you'll see what I mean. Why would this be? No line-height is specified anywhere else in the CSS, so nothing is overriding it. That couldn't be the case anyway because I'm adjusting the line-height up and down within the same selector, so it doesn't make sense that a higher value would apply, but a lower value would get overridden.

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  • PHP DateTime accept multiple formats?

    - by John Smith
    I'm trying to construct a DateTime object with multiple accepted formats. According to the DateTime::createFromFormat docs, the first parameter (format) must be a string. I was wondering if there was a way to createFromFormats. In my case, I want the year for my format to be optional: DateTime::createFromFormat('Y-m-d', $date); DateTime::createFromFormat('m-d', $date); so that a user can input just 'm-d' and the year would be assumed 2013. If I wanted multiple accepted formats, would I have to call createFromFormat each time? Shortest thing for my scenario is: DateTime::createFromFormat('m-d', $date) ?: DateTime::createFromFormat('Y-m-d', $date);

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  • how to visualize (value, count) dataset with thousands data points

    - by user510040
    I have a file with 2 numeric columns: value and count. File may have 5000 rows. I do plot(value, count) to find the shape of distribution. But because there are too many data points the picture is not very clear. Do you know better visualization approach? Probably histograms or barplot with grouping close values on x axis will be the better way to look on data? I cannot figure out the syntax of using histogram or barplot for my case.

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  • CSS center page on screen

    - by Kostronor
    //sorry for the bad formating, i am on my phone... When someone asks how to center a page, then the response is like: margin-left:50%; left:(-1/2 width); I used this code on a site with a width of 1000px,so it comes to screens, where this site does not fit. Now the site gets centered on the smaller screen and gets equaly pushet to left and right. So lets say, our screen is 600px wide: 200px are left 600px are on screen 200px are right You can scroll to the right, but the pixels on the left are unreachable... How can i solve this to control, how much of my site gets dragged to the left in case of smaller screens? This is especially important for mobile phones...

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  • Android search list, String

    - by NightSky
    Hey guys what is the best way to search through my list of objects, they return a few strings, last name and first name for example. Here how i'm currently searching but my search needs to match the entire string which I don't want. The search needs it to match part of the string like our contacts list on our phone and ignore the case. if (searchQ.equalsIgnoreCase(child.first_name)) { addChildToList(child); } Ive tried contains and starts with for example, they did not work. Whats going on? Thanks! Cheers!

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  • Why structs cannot be assigned directly?

    - by becko
    Suppose I have a fully defined struct with tag MyStruct, and suppose that x, y, ..., z are allowed values for its fields. Why is struct MyStruct q = {x,y,..,z}; allowed, but struct MyStruct q; q = {x,y,...,z}; is not allowed? I find this very annoying. In the second case, where I have previously declared q, I need to assign a value to each field, one by one: q.X = x; q.Y = y; ... q.Z = z; where X, Y, ..., Z are the fields of MyStruct. Is there a reason behind this?

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  • Why does SQL Server consider N'????' and N'???' to be equal?

    - by Aidan Ryan
    We are testing our application for Unicode compatibility and have been selecting random characters outside the Latin character set for testing. On both Latin and Japanese-collated systems the following equality is true (U+3422): N'????' = N'???' but the following is not (U+30C1): N'????' = N'???' This was discovered when a test case using the first example (using U+3422) violated a unique index. Do we need to be more selective about the characters we use for testing? Obviously we don't know the semantic meaning of the above comparisons. Would this behavior be obvious to a native speaker?

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  • What is the proper way to test is variable is empty in a batch file, If NOT "%1" == "" GOTO SomeLabe

    - by blak3r
    I need to test if a variable is set or not. I've tried several techniques but they seem to fail whenever %1 is surrounded by quotes such as the case when %1 = "c:\some path with spaces". IF NOT %1 GOTO MyLabel // This is invalid syntax IF "%1" == "" GOTO MyLabel // Works unless %1, otherwise fatally kills bat execution IF %1 == GOTO MyLabel // Gives an unexpected GOTO error. According to this site. These are the support IF syntax types. So, I don't see a way to do it. IF [NOT] ERRORLEVEL number command IF [NOT] string1==string2 command IF [NOT] EXIST filename command

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  • nginx - redirection doesn't work as expected

    - by Luis
    I have a domain listening on both http and https. I want to redirect all the traffic to https except for two specific locations. It works, but only for mydomain.com, not for www.mydomain.com. Here the config: upstream mydomain_rails { server unix:/home/deploy/mydomain/shared/pids/unicorn.sock; } # blog.mydomain.com server { listen 80; server_name blog.mydomain.com; rewrite ^ http://www.mydomain.com/de/blog permanent; } # blog.mydomain.com.br server { listen 80; server_name blog.mydomain.com.br; rewrite ^ http://www.mydomain.com/br/blog permanent; } # www.mydomain.de server { listen 80; server_name mydomain.de www.mydomain.de; rewrite ^ https://www.mydomain.com/de permanent; } # www.mydomain.com.br server { listen 80; server_name mydomain.com.br www.mydomain.com.br; rewrite ^ https://www.mydomain.com/br permanent; } server { listen 80; server_name mydomain.com; rewrite ^ http://www.mydomain.com$request_uri permanent; } ## www.mydomain.com ## Redirect http to https, keep blogs on plain http server { listen 80; server_name www.mydomain.com; location / { # if ($host ~* ^(www\.mydomain\.com)$ ) { rewrite ^/(.*)$ https://www.mydomain.com/$1 permanent; # } # return 444; } # Matches any request starting with '/br/blog' and proxies to the upstream blog instance location ~* /br/blog { proxy_set_header X-Forwarded-For $proxy_add_x_forwarded_for; proxy_set_header Host $http_host; proxy_redirect off; if (!-f $request_filename) { rewrite ^/br/blog$ /; rewrite ^/br/blog/(.*)$ /$1; proxy_pass http://mydomain_blog_br; break; } } # Matches any request starting with '/de/blog' and proxies to the upstream blog instance location ~* /de/blog { proxy_set_header X-Forwarded-For $proxy_add_x_forwarded_for; proxy_set_header Host $http_host; proxy_redirect off; if (!-f $request_filename) { rewrite ^/de/blog$ /; rewrite ^/de/blog/(.*)$ /$1; proxy_pass http://mydomain_blog; break; } } } # www.mydomain.com server { add_header Cache-Control "public, must-revalidate"; server_name mydomain.com www.mydomain.com; listen 443; ssl on; ssl_certificate /etc/ssl/mydomain.com/sslchain.crt; ssl_certificate_key /etc/ssl/mydomain.com/privatekey.key; ## Strict Transport Security (ForceHTTPS), max-age 30d add_header Strict-Transport-Security "max-age=2592000; includeSubdomains"; ## Due SSL encryption, rather to increase the keepalive requests and timeout keepalive_requests 10; keepalive_timeout 60 60; root /home/deploy/mydomain/current/public/; error_log /home/deploy/mydomain/shared/log/nginx.error.log info; access_log /home/deploy/mydomain/shared/log/nginx.access.log main; ## Redirect from non-www to www if ($host = 'mydomain.com' ) { rewrite ^/(.*)$ https://www.mydomain.com/$1 permanent; } ## Caching images for 3 months location ~* \.(ico|css|js|gif|jpe?g|png)\?[0-9]+$ { expires 30d; break; } ## Deny illegal Host headers if ($host !~* ^(mydomain.com|www.mydomain.com)$ ) { return 444; } ## Deny certain User-Agents (case insensitive) if ($http_user_agent ~* (Baiduspider|webalta|Wget|WordPress|youdao|jakarta) ) { return 444; } ## Deny certain Referers (case insensitive) if ($http_referer ~* (dating|diamond|forsale|girl|jewelry|nudit|poker|porn|poweroversoftware|sex|teen|webcam|zippo|zongdo) ) { return 444; } ## Enable maintenance page. The page is copied in during capistrano deployment set $maintenance 0; if (-f $document_root/index.html) { set $maintenance 1; } if ($request_uri ~* (jpg|jpeg|gif|png|js|css)$) { set $maintenance 0; } if ($maintenance) { rewrite ^(.*)$ /index.html last; break; } location /uk { auth_basic "Restricted"; auth_basic_user_file /etc/nginx/htpasswd; root /home/deploy/mydomain/current/public/; try_files $uri @fallback; } # Matches any request starting with '/br/blog' and proxies to the upstream blog instance location ^~ /br/blog { proxy_set_header X-Forwarded-For $proxy_add_x_forwarded_for; proxy_set_header Host $http_host; proxy_redirect off; if (!-f $request_filename) { rewrite ^/br/blog$ /; rewrite ^/br/blog/(.*)$ /$1; proxy_pass http://mydomain_blog_br; break; } } # Matches any request starting with '/de/blog' and proxies to the upstream blog instance location ^~ /de/blog { proxy_set_header X-Forwarded-For $proxy_add_x_forwarded_for; proxy_set_header Host $http_host; proxy_redirect off; if (!-f $request_filename) { rewrite ^/de/blog$ /; rewrite ^/de/blog/(.*)$ /$1; proxy_pass http://mydomain_blog; break; }} # Matches any request starting with '/lp' and proxies to the upstream blog instance location /lp { proxy_set_header X-Forwarded-For $proxy_add_x_forwarded_for; proxy_set_header Host $http_host; proxy_redirect off; rewrite ^/lp(/?.*)$ /$1; proxy_pass http://mydomain_landingpage; break; } #Matches any request, and looks for static files before reverse proxying to the upstream app server socket location / { root /home/deploy/mydomain/current/public/; try_files $uri @fallback; } # Called after the above pattern, if no static file is found location @fallback { proxy_set_header X-Sendfile-Type X-Accel-Redirect; proxy_set_header X-Forwarded-For $proxy_add_x_forwarded_for; proxy_set_header Host $http_host; proxy_redirect off; proxy_pass http://mydomain_rails; } ## All other errors get the generic error page error_page 400 401 402 403 404 405 406 407 408 409 410 411 412 413 414 415 416 417 495 496 497 500 501 502 503 504 505 506 507 /500.html; location /500.html { root /home/deploy/mydomain/current/public/; } } I defined the blog upstream. As said, it works properly for mydomain.com, but not for www.mydomain.com. Any idea?

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  • Does anyone really understand how HFSC scheduling in Linux/BSD works?

    - by Mecki
    I read the original SIGCOMM '97 PostScript paper about HFSC, it is very technically, but I understand the basic concept. Instead of giving a linear service curve (as with pretty much every other scheduling algorithm), you can specify a convex or concave service curve and thus it is possible to decouple bandwidth and delay. However, even though this paper mentions to kind of scheduling algorithms being used (real-time and link-share), it always only mentions ONE curve per scheduling class (the decoupling is done by specifying this curve, only one curve is needed for that). Now HFSC has been implemented for BSD (OpenBSD, FreeBSD, etc.) using the ALTQ scheduling framework and it has been implemented Linux using the TC scheduling framework (part of iproute2). Both implementations added two additional service curves, that were NOT in the original paper! A real-time service curve and an upper-limit service curve. Again, please note that the original paper mentions two scheduling algorithms (real-time and link-share), but in that paper both work with one single service curve. There never have been two independent service curves for either one as you currently find in BSD and Linux. Even worse, some version of ALTQ seems to add an additional queue priority to HSFC (there is no such thing as priority in the original paper either). I found several BSD HowTo's mentioning this priority setting (even though the man page of the latest ALTQ release knows no such parameter for HSFC, so officially it does not even exist). This all makes the HFSC scheduling even more complex than the algorithm described in the original paper and there are tons of tutorials on the Internet that often contradict each other, one claiming the opposite of the other one. This is probably the main reason why nobody really seems to understand how HFSC scheduling really works. Before I can ask my questions, we need a sample setup of some kind. I'll use a very simple one as seen in the image below: Here are some questions I cannot answer because the tutorials contradict each other: What for do I need a real-time curve at all? Assuming A1, A2, B1, B2 are all 128 kbit/s link-share (no real-time curve for either one), then each of those will get 128 kbit/s if the root has 512 kbit/s to distribute (and A and B are both 256 kbit/s of course), right? Why would I additionally give A1 and B1 a real-time curve with 128 kbit/s? What would this be good for? To give those two a higher priority? According to original paper I can give them a higher priority by using a curve, that's what HFSC is all about after all. By giving both classes a curve of [256kbit/s 20ms 128kbit/s] both have twice the priority than A2 and B2 automatically (still only getting 128 kbit/s on average) Does the real-time bandwidth count towards the link-share bandwidth? E.g. if A1 and B1 both only have 64kbit/s real-time and 64kbit/s link-share bandwidth, does that mean once they are served 64kbit/s via real-time, their link-share requirement is satisfied as well (they might get excess bandwidth, but lets ignore that for a second) or does that mean they get another 64 kbit/s via link-share? So does each class has a bandwidth "requirement" of real-time plus link-share? Or does a class only have a higher requirement than the real-time curve if the link-share curve is higher than the real-time curve (current link-share requirement equals specified link-share requirement minus real-time bandwidth already provided to this class)? Is upper limit curve applied to real-time as well, only to link-share, or maybe to both? Some tutorials say one way, some say the other way. Some even claim upper-limit is the maximum for real-time bandwidth + link-share bandwidth? What is the truth? Assuming A2 and B2 are both 128 kbit/s, does it make any difference if A1 and B1 are 128 kbit/s link-share only, or 64 kbit/s real-time and 128 kbit/s link-share, and if so, what difference? If I use the seperate real-time curve to increase priorities of classes, why would I need "curves" at all? Why is not real-time a flat value and link-share also a flat value? Why are both curves? The need for curves is clear in the original paper, because there is only one attribute of that kind per class. But now, having three attributes (real-time, link-share, and upper-limit) what for do I still need curves on each one? Why would I want the curves shape (not average bandwidth, but their slopes) to be different for real-time and link-share traffic? According to the little documentation available, real-time curve values are totally ignored for inner classes (class A and B), they are only applied to leaf classes (A1, A2, B1, B2). If that is true, why does the ALTQ HFSC sample configuration (search for 3.3 Sample configuration) set real-time curves on inner classes and claims that those set the guaranteed rate of those inner classes? Isn't that completely pointless? (note: pshare sets the link-share curve in ALTQ and grate the real-time curve; you can see this in the paragraph above the sample configuration). Some tutorials say the sum of all real-time curves may not be higher than 80% of the line speed, others say it must not be higher than 70% of the line speed. Which one is right or are they maybe both wrong? One tutorial said you shall forget all the theory. No matter how things really work (schedulers and bandwidth distribution), imagine the three curves according to the following "simplified mind model": real-time is the guaranteed bandwidth that this class will always get. link-share is the bandwidth that this class wants to become fully satisfied, but satisfaction cannot be guaranteed. In case there is excess bandwidth, the class might even get offered more bandwidth than necessary to become satisfied, but it may never use more than upper-limit says. For all this to work, the sum of all real-time bandwidths may not be above xx% of the line speed (see question above, the percentage varies). Question: Is this more or less accurate or a total misunderstanding of HSFC? And if assumption above is really accurate, where is prioritization in that model? E.g. every class might have a real-time bandwidth (guaranteed), a link-share bandwidth (not guaranteed) and an maybe an upper-limit, but still some classes have higher priority needs than other classes. In that case I must still prioritize somehow, even among real-time traffic of those classes. Would I prioritize by the slope of the curves? And if so, which curve? The real-time curve? The link-share curve? The upper-limit curve? All of them? Would I give all of them the same slope or each a different one and how to find out the right slope? I still haven't lost hope that there exists at least a hand full of people in this world that really understood HFSC and are able to answer all these questions accurately. And doing so without contradicting each other in the answers would be really nice ;-)

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  • Trouble connecting to vsftpd on ubuntu server

    - by littleK
    I have installed Ubuntu Server 10.10 and I am using it to host a domain that I have. I am trying to set up FTP for the server, but I am running into some problems. I have successfully installed vsFTPd and I have opened up ports 20, 21 on my firewall. In my vsFTPd configuration, I have enabled SSL. Every time I try to connect to my server via FTP, I receive a "Connection Refused" error. I have had a little more success with SSL disabled, however the connection process will time out after the LIST command (but it does accept my authentication). Here is my vsFTPd configuration, the SSL stuff is at the bottom: # Example config file /etc/vsftpd.conf # # The default compiled in settings are fairly paranoid. This sample file # loosens things up a bit, to make the ftp daemon more usable. # Please see vsftpd.conf.5 for all compiled in defaults. # # READ THIS: This example file is NOT an exhaustive list of vsftpd options. # Please read the vsftpd.conf.5 manual page to get a full idea of vsftpd's # capabilities. # # # Run standalone? vsftpd can run either from an inetd or as a standalone # daemon started from an initscript. listen=YES # # Run standalone with IPv6? # Like the listen parameter, except vsftpd will listen on an IPv6 socket # instead of an IPv4 one. This parameter and the listen parameter are mutually # exclusive. #listen_ipv6=YES # # Allow anonymous FTP? (Disabled by default) anonymous_enable=NO # # Uncomment this to allow local users to log in. local_enable=YES # # Uncomment this to enable any form of FTP write command. write_enable=YES # # Default umask for local users is 077. You may wish to change this to 022, # if your users expect that (022 is used by most other ftpd's) #local_umask=022 # # Uncomment this to allow the anonymous FTP user to upload files. This only # has an effect if the above global write enable is activated. Also, you will # obviously need to create a directory writable by the FTP user. #anon_upload_enable=YES # # Uncomment this if you want the anonymous FTP user to be able to create # new directories. #anon_mkdir_write_enable=YES # # Activate directory messages - messages given to remote users when they # go into a certain directory. dirmessage_enable=YES # # If enabled, vsftpd will display directory listings with the time # in your local time zone. The default is to display GMT. The # times returned by the MDTM FTP command are also affected by this # option. use_localtime=YES # # Activate logging of uploads/downloads. xferlog_enable=YES # # Make sure PORT transfer connections originate from port 20 (ftp-data). connect_from_port_20=YES # # If you want, you can arrange for uploaded anonymous files to be owned by # a different user. Note! Using "root" for uploaded files is not # recommended! #chown_uploads=YES #chown_username=whoever # # You may override where the log file goes if you like. The default is shown # below. #xferlog_file=/var/log/vsftpd.log # # If you want, you can have your log file in standard ftpd xferlog format. # Note that the default log file location is /var/log/xferlog in this case. #xferlog_std_format=YES # # You may change the default value for timing out an idle session. #idle_session_timeout=600 # # You may change the default value for timing out a data connection. #data_connection_timeout=120 # # It is recommended that you define on your system a unique user which the # ftp server can use as a totally isolated and unprivileged user. #nopriv_user=ftpsecure # # Enable this and the server will recognise asynchronous ABOR requests. Not # recommended for security (the code is non-trivial). Not enabling it, # however, may confuse older FTP clients. #async_abor_enable=YES # # By default the server will pretend to allow ASCII mode but in fact ignore # the request. Turn on the below options to have the server actually do ASCII # mangling on files when in ASCII mode. # Beware that on some FTP servers, ASCII support allows a denial of service # attack (DoS) via the command "SIZE /big/file" in ASCII mode. vsftpd # predicted this attack and has always been safe, reporting the size of the # raw file. # ASCII mangling is a horrible feature of the protocol. #ascii_upload_enable=YES #ascii_download_enable=YES # # You may fully customise the login banner string: #ftpd_banner=Welcome to blah FTP service. # # You may specify a file of disallowed anonymous e-mail addresses. Apparently # useful for combatting certain DoS attacks. #deny_email_enable=YES # (default follows) #banned_email_file=/etc/vsftpd.banned_emails # # You may restrict local users to their home directories. See the FAQ for # the possible risks in this before using chroot_local_user or # chroot_list_enable below. #chroot_local_user=YES # # You may specify an explicit list of local users to chroot() to their home # directory. If chroot_local_user is YES, then this list becomes a list of # users to NOT chroot(). #chroot_local_user=YES #chroot_list_enable=YES # (default follows) #chroot_list_file=/etc/vsftpd.chroot_list # # You may activate the "-R" option to the builtin ls. This is disabled by # default to avoid remote users being able to cause excessive I/O on large # sites. However, some broken FTP clients such as "ncftp" and "mirror" assume # the presence of the "-R" option, so there is a strong case for enabling it. #ls_recurse_enable=YES # # Debian customization # # Some of vsftpd's settings don't fit the Debian filesystem layout by # default. These settings are more Debian-friendly. # # This option should be the name of a directory which is empty. Also, the # directory should not be writable by the ftp user. This directory is used # as a secure chroot() jail at times vsftpd does not require filesystem # access. secure_chroot_dir=/var/run/vsftpd/empty # # This string is the name of the PAM service vsftpd will use. pam_service_name=vsftpd # # This option specifies the location of the RSA certificate to use for SSL # encrypted connections. rsa_cert_file=/etc/ssl/private/vsftpd.pem # SSL ssl_enable=YES allow_anon_ssl=NO force_local_data_ssl=YES force_local_logins_ssl=YES ssl_tlsv1=YES ssl_sslv2=YES ssl_sslv3=YES Thanks!

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  • Can't Get Virtual Users Setup in VSFTPD -Tried Everything

    - by N.T.
    Have Ubuntu 11.10 with vsftpd installed and working. Can not get virtual users setup at all? Vsftpd will allow main Ubuntu owner account to login, but nothing else? I've followed several tutorials on adding virtual users, but nothing works? I just need to add 2 virtual users and have them be able to upload files to vsftpd Ubuntu computer from other computers on my Lan network. Everywhere I've looked, people just point toward tutorials on adding virtual users, but that just is NOT working. I've been struggling with this for over a week now! PLEASE Help. Thanks. I'll even give a donation if someone can figure this out. here is the vsftpd.conf file I am using. I copied the original, and make a new one, every time I try a tutorial. So far, none have worked. Here is the vsftpd.conf file I'm using. (I hope this helps?) # Example config file /etc/vsftpd.conf # # The default compiled in settings are fairly paranoid. This sample file # loosens things up a bit, to make the ftp daemon more usable. # Please see vsftpd.conf.5 for all compiled in defaults. # # READ THIS: This example file is NOT an exhaustive list of vsftpd options. # Please read the vsftpd.conf.5 manual page to get a full idea of vsftpd's # capabilities. # # # Run standalone? vsftpd can run either from an inetd or as a standalone # daemon started from an initscript. listen=YES # # Run standalone with IPv6? # Like the listen parameter, except vsftpd will listen on an IPv6 socket # instead of an IPv4 one. This parameter and the listen parameter are mutually # exclusive. #listen_ipv6=YES # # Allow anonymous FTP? (Disabled by default) anonymous_enable=YES # # Uncomment this to allow local users to log in. local_enable=YES # # Uncomment this to enable any form of FTP write command. write_enable=YES # # Default umask for local users is 077. You may wish to change this to 022, # if your users expect that (022 is used by most other ftpd's) local_umask=022 # # Uncomment this to allow the anonymous FTP user to upload files. This only # has an effect if the above global write enable is activated. Also, you will # obviously need to create a directory writable by the FTP user. #anon_upload_enable=YES # # Uncomment this if you want the anonymous FTP user to be able to create # new directories. anon_mkdir_write_enable=YES # # Activate directory messages - messages given to remote users when they # go into a certain directory. dirmessage_enable=YES # # If enabled, vsftpd will display directory listings with the time # in your local time zone. The default is to display GMT. The # times returned by the MDTM FTP command are also affected by this # option. use_localtime=YES # # Activate logging of uploads/downloads. xferlog_enable=YES # # Make sure PORT transfer connections originate from port 20 (ftp-data). connect_from_port_20=YES # # If you want, you can arrange for uploaded anonymous files to be owned by # a different user. Note! Using "root" for uploaded files is not # recommended! #chown_uploads=YES #chown_username=whoever # # You may override where the log file goes if you like. The default is shown # below. #xferlog_file=/var/log/vsftpd.log # # If you want, you can have your log file in standard ftpd xferlog format. # Note that the default log file location is /var/log/xferlog in this case. xferlog_std_format=YES # # You may change the default value for timing out an idle session. #idle_session_timeout=600 # # You may change the default value for timing out a data connection. #data_connection_timeout=120 # # It is recommended that you define on your system a unique user which the # ftp server can use as a totally isolated and unprivileged user. #nopriv_user=ftpsecure # # Enable this and the server will recognise asynchronous ABOR requests. Not # recommended for security (the code is non-trivial). Not enabling it, # however, may confuse older FTP clients. #async_abor_enable=YES # # By default the server will pretend to allow ASCII mode but in fact ignore # the request. Turn on the below options to have the server actually do ASCII # mangling on files when in ASCII mode. # Beware that on some FTP servers, ASCII support allows a denial of service # attack (DoS) via the command "SIZE /big/file" in ASCII mode. vsftpd # predicted this attack and has always been safe, reporting the size of the # raw file. # ASCII mangling is a horrible feature of the protocol. #ascii_upload_enable=YES #ascii_download_enable=YES # # You may fully customise the login banner string: ftpd_banner=Welcome to Sage FTP service. # # You may specify a file of disallowed anonymous e-mail addresses. Apparently # useful for combatting certain DoS attacks. #deny_email_enable=YES # (default follows) #banned_email_file=/etc/vsftpd.banned_emails # # You may restrict local users to their home directories. See the FAQ for # the possible risks in this before using chroot_local_user or # chroot_list_enable below. chroot_local_user=YES # # You may specify an explicit list of local users to chroot() to their home # directory. If chroot_local_user is YES, then this list becomes a list of # users to NOT chroot(). #chroot_local_user=YES #chroot_list_enable=YES # (default follows) #chroot_list_file=/etc/vsftpd.chroot_list # # You may activate the "-R" option to the builtin ls. This is disabled by # default to avoid remote users being able to cause excessive I/O on large # sites. However, some broken FTP clients such as "ncftp" and "mirror" assume # the presence of the "-R" option, so there is a strong case for enabling it. #ls_recurse_enable=YES # # Debian customization # # Some of vsftpd's settings don't fit the Debian filesystem layout by # default. These settings are more Debian-friendly. # # This option should be the name of a directory which is empty. Also, the # directory should not be writable by the ftp user. This directory is used # as a secure chroot() jail at times vsftpd does not require filesystem # access. secure_chroot_dir=/var/run/vsftpd/empty # # This string is the name of the PAM service vsftpd will use. pam_service_name=vsftpd local_root=/media/FilesDrive # # This option specifies the location of the RSA certificate to use for SSL # encrypted connections. rsa_cert_file=/etc/ssl/private/vsftpd.pem

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  • Does anyone really understand how HFSC scheduling in Linux/BSD works?

    - by Mecki
    I read the original SIGCOMM '97 PostScript paper about HFSC, it is very technically, but I understand the basic concept. Instead of giving a linear service curve (as with pretty much every other scheduling algorithm), you can specify a convex or concave service curve and thus it is possible to decouple bandwidth and delay. However, even though this paper mentions to kind of scheduling algorithms being used (real-time and link-share), it always only mentions ONE curve per scheduling class (the decoupling is done by specifying this curve, only one curve is needed for that). Now HFSC has been implemented for BSD (OpenBSD, FreeBSD, etc.) using the ALTQ scheduling framework and it has been implemented Linux using the TC scheduling framework (part of iproute2). Both implementations added two additional service curves, that were NOT in the original paper! A real-time service curve and an upper-limit service curve. Again, please note that the original paper mentions two scheduling algorithms (real-time and link-share), but in that paper both work with one single service curve. There never have been two independent service curves for either one as you currently find in BSD and Linux. Even worse, some version of ALTQ seems to add an additional queue priority to HSFC (there is no such thing as priority in the original paper either). I found several BSD HowTo's mentioning this priority setting (even though the man page of the latest ALTQ release knows no such parameter for HSFC, so officially it does not even exist). This all makes the HFSC scheduling even more complex than the algorithm described in the original paper and there are tons of tutorials on the Internet that often contradict each other, one claiming the opposite of the other one. This is probably the main reason why nobody really seems to understand how HFSC scheduling really works. Before I can ask my questions, we need a sample setup of some kind. I'll use a very simple one as seen in the image below: Here are some questions I cannot answer because the tutorials contradict each other: What for do I need a real-time curve at all? Assuming A1, A2, B1, B2 are all 128 kbit/s link-share (no real-time curve for either one), then each of those will get 128 kbit/s if the root has 512 kbit/s to distribute (and A and B are both 256 kbit/s of course), right? Why would I additionally give A1 and B1 a real-time curve with 128 kbit/s? What would this be good for? To give those two a higher priority? According to original paper I can give them a higher priority by using a curve, that's what HFSC is all about after all. By giving both classes a curve of [256kbit/s 20ms 128kbit/s] both have twice the priority than A2 and B2 automatically (still only getting 128 kbit/s on average) Does the real-time bandwidth count towards the link-share bandwidth? E.g. if A1 and B1 both only have 64kbit/s real-time and 64kbit/s link-share bandwidth, does that mean once they are served 64kbit/s via real-time, their link-share requirement is satisfied as well (they might get excess bandwidth, but lets ignore that for a second) or does that mean they get another 64 kbit/s via link-share? So does each class has a bandwidth "requirement" of real-time plus link-share? Or does a class only have a higher requirement than the real-time curve if the link-share curve is higher than the real-time curve (current link-share requirement equals specified link-share requirement minus real-time bandwidth already provided to this class)? Is upper limit curve applied to real-time as well, only to link-share, or maybe to both? Some tutorials say one way, some say the other way. Some even claim upper-limit is the maximum for real-time bandwidth + link-share bandwidth? What is the truth? Assuming A2 and B2 are both 128 kbit/s, does it make any difference if A1 and B1 are 128 kbit/s link-share only, or 64 kbit/s real-time and 128 kbit/s link-share, and if so, what difference? If I use the seperate real-time curve to increase priorities of classes, why would I need "curves" at all? Why is not real-time a flat value and link-share also a flat value? Why are both curves? The need for curves is clear in the original paper, because there is only one attribute of that kind per class. But now, having three attributes (real-time, link-share, and upper-limit) what for do I still need curves on each one? Why would I want the curves shape (not average bandwidth, but their slopes) to be different for real-time and link-share traffic? According to the little documentation available, real-time curve values are totally ignored for inner classes (class A and B), they are only applied to leaf classes (A1, A2, B1, B2). If that is true, why does the ALTQ HFSC sample configuration (search for 3.3 Sample configuration) set real-time curves on inner classes and claims that those set the guaranteed rate of those inner classes? Isn't that completely pointless? (note: pshare sets the link-share curve in ALTQ and grate the real-time curve; you can see this in the paragraph above the sample configuration). Some tutorials say the sum of all real-time curves may not be higher than 80% of the line speed, others say it must not be higher than 70% of the line speed. Which one is right or are they maybe both wrong? One tutorial said you shall forget all the theory. No matter how things really work (schedulers and bandwidth distribution), imagine the three curves according to the following "simplified mind model": real-time is the guaranteed bandwidth that this class will always get. link-share is the bandwidth that this class wants to become fully satisfied, but satisfaction cannot be guaranteed. In case there is excess bandwidth, the class might even get offered more bandwidth than necessary to become satisfied, but it may never use more than upper-limit says. For all this to work, the sum of all real-time bandwidths may not be above xx% of the line speed (see question above, the percentage varies). Question: Is this more or less accurate or a total misunderstanding of HSFC? And if assumption above is really accurate, where is prioritization in that model? E.g. every class might have a real-time bandwidth (guaranteed), a link-share bandwidth (not guaranteed) and an maybe an upper-limit, but still some classes have higher priority needs than other classes. In that case I must still prioritize somehow, even among real-time traffic of those classes. Would I prioritize by the slope of the curves? And if so, which curve? The real-time curve? The link-share curve? The upper-limit curve? All of them? Would I give all of them the same slope or each a different one and how to find out the right slope? I still haven't lost hope that there exists at least a hand full of people in this world that really understood HFSC and are able to answer all these questions accurately. And doing so without contradicting each other in the answers would be really nice ;-)

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  • Why are my hard drives failing?

    - by WishCow
    I have a small Ubuntu server running at home, with 2 HDDs. There are two software raids (raid1) on the disks, managed by mdadm, which I believe is irrelevant, but mentioning it anyway. Both of the HDDs are Western Digital, and have been used for around 2 years, when one of them started making clicking noises, and died. I figured that maybe it's natural after 2 years, so I bought a new one, and resynced the raid arrays. After about a month, the other drive also died. I didn't get suspicious, since both drives have been bought at the same time, it's not that surprising to see both of them near each other, so I bought another one. So far, 2 old drives failed, and 2 brand new in the system. After one month, one of the new drives died. This is when it started getting suspicious. Since the PC was put together from some really old parts (think AthlonXP), I figured that maybe the motherboard's SATA controller is the culprit. Of course you can't switch parts easily in an old PC like this, so I bought a whole system, new MB, new CPU, new RAM. Took the just failed drive back, since it was under warranty, and got it replaced. So it is up to 2 failed drives from the old ones, and 1 failed drive from the new ones. No problems, for 1 month. After that errors were creeping up again in /var/log/messages, and mdadm was reporting raid array failures. I started tearing my hair out. Everything is new in the system, it's up to the third brand new HDD, it's simply not possible that all of the new drives that I bought were faulty. Let's see what is still common... the cables. Okay, long shot, let's replace the SATA cables. Take HDD back, smile to the guy at the counter and say that I'm really unlucky. He replaces the HDD. I come home, one month passes and one of HDDs fails, again. I'm not joking. Two of the brand new HDDs have failed. Maybe it's a bug in the OS. Let's see what the manufacturer's testing tool says. Download testing tool, burn it to a CD, reboot, leave HDD testing overnight. Test says that the drive is faulty, and I should back up everything, if I still can. I don't know what's happening, but it does not look like a software problem, something is definitely thrashing the HDDs. I should mention now, that the whole system is in a shoebox. Since there are a load of "build your own ikea case" stuff, I thought there shouldn't be any problems throwing the thing in a box, and stuffing it away somewhere. The box is well ventilated, but I thought that just maybe the drives were overheating. There is no other possible answer to this. So I took the HDD back, and got it replaced (for the 3rd time), and bought HDD coolers. And just now, I have heard the sound of doom. click click whizzzzzzzzz. SSH into the box: You have new mail! mail r 1 DegradedArrayEvent on /dev/md0 ... dmesg output: [47128.000051] ata3: lost interrupt (Status 0x50) [47128.000097] end_request: I/O error, dev sda, sector 58588863 [47128.000134] md: super_written gets error=-5, uptodate=0 [48043.976054] ata3: lost interrupt (Status 0x50) [48043.976086] ata3.00: exception Emask 0x0 SAct 0x0 SErr 0x0 action 0x6 frozen [48043.976132] ata3.00: cmd c8/00:18:bf:40:52/00:00:00:00:00/e1 tag 0 dma 12288 in [48043.976135] res 40/00:00:00:4f:c2/00:00:00:00:00/00 Emask 0x4 (timeout) [48043.976208] ata3.00: status: { DRDY } [48043.976241] ata3: soft resetting link [48044.148446] ata3.00: configured for UDMA/133 [48044.148457] ata3.00: device reported invalid CHS sector 0 [48044.148477] ata3: EH complete Recap: No possibility of overheating 6 drives have failed, 4 of those have been brand new. I'm not sure now that the original two have been faulty, or suffered the same thing that the new ones. There is nothing common in the system, apart from the OS which is Ubuntu Karmic now (started with Jaunty). New MB, new CPU, new RAM, new SATA cables. No, the little holes on the HDD are not covered I'm crying. Really. I don't have the face to return to the store now, it's not possible for 4 drives to fail under 4 months. A few ideas that I have been thinking: Is it possible that I fuck up something when I partition and resync the drives? Can it be so bad that it physicaly wrecks the drive? (since the vendor supplied tool says that the drive is damaged) I do the partitoning with fdisk, and use the same block size for the raid1 partitions (I check the exact block sizes with fdisk -lu) Is it possible that the linux kernel or mdadm, or something is not compatible with this exact brand of HDDs, and thrashes them? Is it possible that it may be the shoebox? Try placing it somewhere else? It's under a shelf now, so humidity is not a problem either. Is it possible that a normal PC case will solve my problem (I'm going to shoot myself then)? I will get a picture tomorrow. Am I just simply cursed? Any help or speculation is greatly appreciated. Edit: The power strip is guarded against overvoltage. Edit2: I have moved inbetween these 4 months, so the possibility of the cause being "dirty" electricity in both places, is very low. Edit3: I have checked the voltages in the BIOS (couldn't borrow a multimeter), and they are all seem correct, the biggest discrepancy is in the 12V, because it's supplying 11.3. Should I be worried about that? Edit4: I put my desktop PC's PSU into the server. The BIOS reported much more accurate voltage readings, and also it has successfully rebuilt the raid1 array, which took some 3-4 hours, so I feel a little positive now. Will get a new PSU tomorrow to test with that. Also, attaching the picture about the box: (disregard the 3rd drive)

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  • e2fsck extremely slow, although enough memory exists

    - by kaefert
    I've got this external USB-Disk: kaefert@blechmobil:~$ lsusb -s 2:3 Bus 002 Device 003: ID 0bc2:3320 Seagate RSS LLC As can be seen in this dmesg output, there is some problem that prevents that disk from beeing mounted: kaefert@blechmobil:~$ dmesg ... [ 113.084079] usb 2-1: new high-speed USB device number 3 using ehci_hcd [ 113.217783] usb 2-1: New USB device found, idVendor=0bc2, idProduct=3320 [ 113.217787] usb 2-1: New USB device strings: Mfr=2, Product=3, SerialNumber=1 [ 113.217790] usb 2-1: Product: Expansion Desk [ 113.217792] usb 2-1: Manufacturer: Seagate [ 113.217794] usb 2-1: SerialNumber: NA4J4N6K [ 113.435404] usbcore: registered new interface driver uas [ 113.455315] Initializing USB Mass Storage driver... [ 113.468051] scsi5 : usb-storage 2-1:1.0 [ 113.468180] usbcore: registered new interface driver usb-storage [ 113.468182] USB Mass Storage support registered. [ 114.473105] scsi 5:0:0:0: Direct-Access Seagate Expansion Desk 070B PQ: 0 ANSI: 6 [ 114.474342] sd 5:0:0:0: [sdb] 732566645 4096-byte logical blocks: (3.00 TB/2.72 TiB) [ 114.475089] sd 5:0:0:0: [sdb] Write Protect is off [ 114.475092] sd 5:0:0:0: [sdb] Mode Sense: 43 00 00 00 [ 114.475959] sd 5:0:0:0: [sdb] Write cache: enabled, read cache: enabled, doesn't support DPO or FUA [ 114.477093] sd 5:0:0:0: [sdb] 732566645 4096-byte logical blocks: (3.00 TB/2.72 TiB) [ 114.501649] sdb: sdb1 [ 114.502717] sd 5:0:0:0: [sdb] 732566645 4096-byte logical blocks: (3.00 TB/2.72 TiB) [ 114.504354] sd 5:0:0:0: [sdb] Attached SCSI disk [ 116.804408] EXT4-fs (sdb1): ext4_check_descriptors: Checksum for group 3976 failed (47397!=61519) [ 116.804413] EXT4-fs (sdb1): group descriptors corrupted! ... So I went and fired up my favorite partition manager - gparted, and told it to verify and repair the partition sdb1. This made gparted call e2fsck (version 1.42.4 (12-Jun-2012)) e2fsck -f -y -v /dev/sdb1 Although gparted called e2fsck with the "-v" option, sadly it doesn't show me the output of my e2fsck process (bugreport https://bugzilla.gnome.org/show_bug.cgi?id=467925 ) I started this whole thing on Sunday (2012-11-04_2200) evening, so about 48 hours ago, this is what htop says about it now (2012-11-06-1900): PID USER PRI NI VIRT RES SHR S CPU% MEM% TIME+ Command 3704 root 39 19 1560M 1166M 768 R 98.0 19.5 42h56:43 e2fsck -f -y -v /dev/sdb1 Now I found a few posts on the internet that discuss e2fsck running slow, for example: http://gparted-forum.surf4.info/viewtopic.php?id=13613 where they write that its a good idea to see if the disk is just that slow because maybe its damaged, and I think these outputs tell me that this is not the case in my case: kaefert@blechmobil:~$ sudo hdparm -tT /dev/sdb /dev/sdb: Timing cached reads: 3562 MB in 2.00 seconds = 1783.29 MB/sec Timing buffered disk reads: 82 MB in 3.01 seconds = 27.26 MB/sec kaefert@blechmobil:~$ sudo hdparm /dev/sdb /dev/sdb: multcount = 0 (off) readonly = 0 (off) readahead = 256 (on) geometry = 364801/255/63, sectors = 5860533160, start = 0 However, although I can read quickly from that disk, this disk speed doesn't seem to be used by e2fsck, considering tools like gkrellm or iotop or this: kaefert@blechmobil:~$ iostat -x Linux 3.2.0-2-amd64 (blechmobil) 2012-11-06 _x86_64_ (2 CPU) avg-cpu: %user %nice %system %iowait %steal %idle 14,24 47,81 14,63 0,95 0,00 22,37 Device: rrqm/s wrqm/s r/s w/s rkB/s wkB/s avgrq-sz avgqu-sz await r_await w_await svctm %util sda 0,59 8,29 2,42 5,14 43,17 160,17 53,75 0,30 39,80 8,72 54,42 3,95 2,99 sdb 137,54 5,48 9,23 0,20 587,07 22,73 129,35 0,07 7,70 7,51 16,18 2,17 2,04 Now I researched a little bit on how to find out what e2fsck is doing with all that processor time, and I found the tool strace, which gives me this: kaefert@blechmobil:~$ sudo strace -p3704 lseek(4, 41026998272, SEEK_SET) = 41026998272 write(4, "\212\354K[_\361\3nl\212\245\352\255jR\303\354\312Yv\334p\253r\217\265\3567\325\257\3766"..., 4096) = 4096 lseek(4, 48404766720, SEEK_SET) = 48404766720 read(4, "\7t\260\366\346\337\304\210\33\267j\35\377'\31f\372\252\ffU\317.y\211\360\36\240c\30`\34"..., 4096) = 4096 lseek(4, 41027002368, SEEK_SET) = 41027002368 write(4, "\232]7Ws\321\352\t\1@[+5\263\334\276{\343zZx\352\21\316`1\271[\202\350R`"..., 4096) = 4096 lseek(4, 48404770816, SEEK_SET) = 48404770816 read(4, "\17\362r\230\327\25\346//\210H\v\311\3237\323K\304\306\361a\223\311\324\272?\213\tq \370\24"..., 4096) = 4096 lseek(4, 41027006464, SEEK_SET) = 41027006464 write(4, "\367yy>x\216?=\324Z\305\351\376&\25\244\210\271\22\306}\276\237\370(\214\205G\262\360\257#"..., 4096) = 4096 lseek(4, 48404774912, SEEK_SET) = 48404774912 read(4, "\365\25\0\21|T\0\21}3t_\272\373\222k\r\177\303\1\201\261\221$\261B\232\3142\21U\316"..., 4096) = 4096 ^CProcess 3704 detached around 16 of these lines every second, so 4 read and 4 write operations every second, which I don't consider to be a lot.. And finally, my question: Will this process ever finish? If those numbers from fseek (48404774912) represent bytes, that would be something like 45 gigabytes, with this beeing a 3 terrabyte disk, which would give me 134 days to go, if the speed stays constant, and e2fsck scans the disk like this completly and only once. Do you have some advice for me? I have most of the data on that disk elsewhere, but I've put a lot of hours into sorting and merging it to this disk, so I would prefer to getting this disk up and running again, without formatting it anew. I don't think that the hardware is damaged since the disk is only a few months and since I can't see any I/O errors in the dmesg output. UPDATE: I just looked at the strace output again (2012-11-06_2300), now it looks like this: lseek(4, 1419860611072, SEEK_SET) = 1419860611072 read(4, "3#\f\2447\335\0\22A\355\374\276j\204'\207|\217V|\23\245[\7VP\251\242\276\207\317:"..., 4096) = 4096 lseek(4, 43018145792, SEEK_SET) = 43018145792 write(4, "]\206\231\342Y\204-2I\362\242\344\6R\205\361\324\177\265\317C\334V\324\260\334\275t=\10F."..., 4096) = 4096 lseek(4, 1419860615168, SEEK_SET) = 1419860615168 read(4, "\262\305\314Y\367\37x\326\245\226\226\320N\333$s\34\204\311\222\7\315\236\336\300TK\337\264\236\211n"..., 4096) = 4096 lseek(4, 43018149888, SEEK_SET) = 43018149888 write(4, "\271\224m\311\224\25!I\376\16;\377\0\223H\25Yd\201Y\342\r\203\271\24eG<\202{\373V"..., 4096) = 4096 lseek(4, 1419860619264, SEEK_SET) = 1419860619264 read(4, ";d\360\177\n\346\253\210\222|\250\352T\335M\33\260\320\261\7g\222P\344H?t\240\20\2548\310"..., 4096) = 4096 lseek(4, 43018153984, SEEK_SET) = 43018153984 write(4, "\360\252j\317\310\251G\227\335{\214`\341\267\31Y\202\360\v\374\307oq\3063\217Z\223\313\36D\211"..., 4096) = 4096 So the numbers in the lseek lines before the reads, like 1419860619264 are already a lot bigger, standing for 1.29 terabytes if those numbers are bytes, so it doesn't seem to be a linear progress on a big scale, maybe there are only some areas that need work, that have big gaps in between them. UPDATE2: Okey, big disappointment, the numbers are back to very small again (2012-11-07_0720) lseek(4, 52174548992, SEEK_SET) = 52174548992 read(4, "\374\312\22\\\325\215\213\23\0357U\222\246\370v^f(\312|f\212\362\343\375\373\342\4\204mU6"..., 4096) = 4096 lseek(4, 46603526144, SEEK_SET) = 46603526144 write(4, "\370\261\223\227\23?\4\4\217\264\320_Am\246CQ\313^\203U\253\274\204\277\2564n\227\177\267\343"..., 4096) = 4096 so either e2fsck goes over the data multiple times, or it just hops back and forth multiple times. Or my assumption that those numbers are bytes is wrong. UPDATE3: Since it's mentioned here http://forums.fedoraforum.org/showthread.php?t=282125&page=2 that you can testisk while e2fsck is running, i tried that, though not with a lot of success. When asking testdisk to display the data of my partition, this is what I get: TestDisk 6.13, Data Recovery Utility, November 2011 Christophe GRENIER <[email protected]> http://www.cgsecurity.org 1 P Linux 0 4 5 45600 40 8 732566272 Can't open filesystem. Filesystem seems damaged. And this is what strace currently gives me (2012-11-07_1030) lseek(4, 212460343296, SEEK_SET) = 212460343296 read(4, "\315Mb\265v\377Gn \24\f\205EHh\2349~\330\273\203\3375\206\10\r3=W\210\372\352"..., 4096) = 4096 lseek(4, 47347830784, SEEK_SET) = 47347830784 write(4, "]\204\223\300I\357\4\26\33+\243\312G\230\250\371*m2U\t_\215\265J \252\342Pm\360D"..., 4096) = 4096 (times are in CET)

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  • Finding a Relative Path in .NET

    - by Rick Strahl
    Here’s a nice and simple path utility that I’ve needed in a number of applications: I need to find a relative path based on a base path. So if I’m working in a folder called c:\temp\templates\ and I want to find a relative path for c:\temp\templates\subdir\test.txt I want to receive back subdir\test.txt. Or if I pass c:\ I want to get back ..\..\ – in other words always return a non-hardcoded path based on some other known directory. I’ve had a routine in my library that does this via some lengthy string parsing routines, but ran into some Uri processing today that made me realize that this code could be greatly simplified by using the System.Uri class instead. Here’s the simple static method: /// <summary> /// Returns a relative path string from a full path based on a base path /// provided. /// </summary> /// <param name="fullPath">The path to convert. Can be either a file or a directory</param> /// <param name="basePath">The base path on which relative processing is based. Should be a directory.</param> /// <returns> /// String of the relative path. /// /// Examples of returned values: /// test.txt, ..\test.txt, ..\..\..\test.txt, ., .., subdir\test.txt /// </returns> public static string GetRelativePath(string fullPath, string basePath ) { // ForceBasePath to a path if (!basePath.EndsWith("\\")) basePath += "\\"; Uri baseUri = new Uri(basePath); Uri fullUri = new Uri(fullPath); Uri relativeUri = baseUri.MakeRelativeUri(fullUri); // Uri's use forward slashes so convert back to backward slashes return relativeUri.ToString().Replace("/", "\\"); } You can then call it like this: string relPath = FileUtils.GetRelativePath("c:\temp\templates","c:\temp\templates\subdir\test.txt") It’s not exactly rocket science but it’s useful in many scenarios where you’re working with files based on an application base directory. Right now I’m working on a templating solution (using the Razor Engine) where templates live in a base directory and are supplied as relative paths to that base directory. Resolving these relative paths both ways is important in order to properly check for existance of files and their change status in this case. Not the kind of thing you use every day, but useful to remember.© Rick Strahl, West Wind Technologies, 2005-2010Posted in .NET  CSharp  

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  • Microsoft and jQuery

    - by Rick Strahl
    The jQuery JavaScript library has been steadily getting more popular and with recent developments from Microsoft, jQuery is also getting ever more exposure on the ASP.NET platform including now directly from Microsoft. jQuery is a light weight, open source DOM manipulation library for JavaScript that has changed how many developers think about JavaScript. You can download it and find more information on jQuery on www.jquery.com. For me jQuery has had a huge impact on how I develop Web applications and was probably the main reason I went from dreading to do JavaScript development to actually looking forward to implementing client side JavaScript functionality. It has also had a profound impact on my JavaScript skill level for me by seeing how the library accomplishes things (and often reviewing the terse but excellent source code). jQuery made an uncomfortable development platform (JavaScript + DOM) a joy to work on. Although jQuery is by no means the only JavaScript library out there, its ease of use, small size, huge community of plug-ins and pure usefulness has made it easily the most popular JavaScript library available today. As a long time jQuery user, I’ve been excited to see the developments from Microsoft that are bringing jQuery to more ASP.NET developers and providing more integration with jQuery for ASP.NET’s core features rather than relying on the ASP.NET AJAX library. Microsoft and jQuery – making Friends jQuery is an open source project but in the last couple of years Microsoft has really thrown its weight behind supporting this open source library as a supported component on the Microsoft platform. When I say supported I literally mean supported: Microsoft now offers actual tech support for jQuery as part of their Product Support Services (PSS) as jQuery integration has become part of several of the ASP.NET toolkits and ships in several of the default Web project templates in Visual Studio 2010. The ASP.NET MVC 3 framework (still in Beta) also uses jQuery for a variety of client side support features including client side validation and we can look forward toward more integration of client side functionality via jQuery in both MVC and WebForms in the future. In other words jQuery is becoming an optional but included component of the ASP.NET platform. PSS support means that support staff will answer jQuery related support questions as part of any support incidents related to ASP.NET which provides some piece of mind to some corporate development shops that require end to end support from Microsoft. In addition to including jQuery and supporting it, Microsoft has also been getting involved in providing development resources for extending jQuery’s functionality via plug-ins. Microsoft’s last version of the Microsoft Ajax Library – which is the successor to the native ASP.NET AJAX Library – included some really cool functionality for client templates, databinding and localization. As it turns out Microsoft has rebuilt most of that functionality using jQuery as the base API and provided jQuery plug-ins of these components. Very recently these three plug-ins were submitted and have been approved for inclusion in the official jQuery plug-in repository and been taken over by the jQuery team for further improvements and maintenance. Even more surprising: The jQuery-templates component has actually been approved for inclusion in the next major update of the jQuery core in jQuery V1.5, which means it will become a native feature that doesn’t require additional script files to be loaded. Imagine this – an open source contribution from Microsoft that has been accepted into a major open source project for a core feature improvement. Microsoft has come a long way indeed! What the Microsoft Involvement with jQuery means to you For Microsoft jQuery support is a strategic decision that affects their direction in client side development, but nothing stopped you from using jQuery in your applications prior to Microsoft’s official backing and in fact a large chunk of developers did so readily prior to Microsoft’s announcement. Official support from Microsoft brings a few benefits to developers however. jQuery support in Visual Studio 2010 means built-in support for jQuery IntelliSense, automatically added jQuery scripts in many projects types and a common base for client side functionality that actually uses what most developers are already using. If you have already been using jQuery and were worried about straying from the Microsoft line and their internal Microsoft Ajax Library – worry no more. With official support and the change in direction towards jQuery Microsoft is now following along what most in the ASP.NET community had already been doing by using jQuery, which is likely the reason for Microsoft’s shift in direction in the first place. ASP.NET AJAX and the Microsoft AJAX Library weren’t bad technology – there was tons of useful functionality buried in these libraries. However, these libraries never got off the ground, mainly because early incarnations were squarely aimed at control/component developers rather than application developers. For all the functionality that these controls provided for control developers they lacked in useful and easily usable application developer functionality that was easily accessible in day to day client side development. The result was that even though Microsoft shipped support for these tools in the box (in .NET 3.5 and 4.0), other than for the internal support in ASP.NET for things like the UpdatePanel and the ASP.NET AJAX Control Toolkit as well as some third party vendors, the Microsoft client libraries were largely ignored by the developer community opening the door for other client side solutions. Microsoft seems to be acknowledging developer choice in this case: Many more developers were going down the jQuery path rather than using the Microsoft built libraries and there seems to be little sense in continuing development of a technology that largely goes unused by the majority of developers. Kudos for Microsoft for recognizing this and gracefully changing directions. Note that even though there will be no further development in the Microsoft client libraries they will continue to be supported so if you’re using them in your applications there’s no reason to start running for the exit in a panic and start re-writing everything with jQuery. Although that might be a reasonable choice in some cases, jQuery and the Microsoft libraries work well side by side so that you can leave existing solutions untouched even as you enhance them with jQuery. The Microsoft jQuery Plug-ins – Solid Core Features One of the most interesting developments in Microsoft’s embracing of jQuery is that Microsoft has started contributing to jQuery via standard mechanism set for jQuery developers: By submitting plug-ins. Microsoft took some of the nicest new features of the unpublished Microsoft Ajax Client Library and re-wrote these components for jQuery and then submitted them as plug-ins to the jQuery plug-in repository. Accepted plug-ins get taken over by the jQuery team and that’s exactly what happened with the three plug-ins submitted by Microsoft with the templating plug-in even getting slated to be published as part of the jQuery core in the next major release (1.5). The following plug-ins are provided by Microsoft: jQuery Templates – a client side template rendering engine jQuery Data Link – a client side databinder that can synchronize changes without code jQuery Globalization – provides formatting and conversion features for dates and numbers The first two are ports of functionality that was slated for the Microsoft Ajax Library while functionality for the globalization library provides functionality that was already found in the original ASP.NET AJAX library. To me all three plug-ins address a pressing need in client side applications and provide functionality I’ve previously used in other incarnations, but with more complete implementations. Let’s take a close look at these plug-ins. jQuery Templates http://api.jquery.com/category/plugins/templates/ Client side templating is a key component for building rich JavaScript applications in the browser. Templating on the client lets you avoid from manually creating markup by creating DOM nodes and injecting them individually into the document via code. Rather you can create markup templates – similar to the way you create classic ASP server markup – and merge data into these templates to render HTML which you can then inject into the document or replace existing content with. Output from templates are rendered as a jQuery matched set and can then be easily inserted into the document as needed. Templating is key to minimize client side code and reduce repeated code for rendering logic. Instead a single template can be used in many places for updating and adding content to existing pages. Further if you build pure AJAX interfaces that rely entirely on client rendering of the initial page content, templates allow you to a use a single markup template to handle all rendering of each specific HTML section/element. I’ve used a number of different client rendering template engines with jQuery in the past including jTemplates (a PHP style templating engine) and a modified version of John Resig’s MicroTemplating engine which I built into my own set of libraries because it’s such a commonly used feature in my client side applications. jQuery templates adds a much richer templating model that allows for sub-templates and access to the data items. Like John Resig’s original Micro Template engine, the core basics of the templating engine create JavaScript code which means that templates can include JavaScript code. To give you a basic idea of how templates work imagine I have an application that downloads a set of stock quotes based on a symbol list then displays them in the document. To do this you can create an ‘item’ template that describes how each of the quotes is renderd as a template inside of the document: <script id="stockTemplate" type="text/x-jquery-tmpl"> <div id="divStockQuote" class="errordisplay" style="width: 500px;"> <div class="label">Company:</div><div><b>${Company}(${Symbol})</b></div> <div class="label">Last Price:</div><div>${LastPrice}</div> <div class="label">Net Change:</div><div> {{if NetChange > 0}} <b style="color:green" >${NetChange}</b> {{else}} <b style="color:red" >${NetChange}</b> {{/if}} </div> <div class="label">Last Update:</div><div>${LastQuoteTimeString}</div> </div> </script> The ‘template’ is little more than HTML with some markup expressions inside of it that define the template language. Notice the embedded ${} expressions which reference data from the quote objects returned from an AJAX call on the server. You can embed any JavaScript or value expression in these template expressions. There are also a number of structural commands like {{if}} and {{each}} that provide for rudimentary logic inside of your templates as well as commands ({{tmpl}} and {{wrap}}) for nesting templates. You can find more about the full set of markup expressions available in the documentation. To load up this data you can use code like the following: <script type="text/javascript"> //var Proxy = new ServiceProxy("../PageMethods/PageMethodsService.asmx/"); $(document).ready(function () { $("#btnGetQuotes").click(GetQuotes); }); function GetQuotes() { var symbols = $("#txtSymbols").val().split(","); $.ajax({ url: "../PageMethods/PageMethodsService.asmx/GetStockQuotes", data: JSON.stringify({ symbols: symbols }), // parameter map type: "POST", // data has to be POSTed contentType: "application/json", timeout: 10000, dataType: "json", success: function (result) { var quotes = result.d; var jEl = $("#stockTemplate").tmpl(quotes); $("#quoteDisplay").empty().append(jEl); }, error: function (xhr, status) { alert(status + "\r\n" + xhr.responseText); } }); }; </script> In this case an ASMX AJAX service is called to retrieve the stock quotes. The service returns an array of quote objects. The result is returned as an object with the .d property (in Microsoft service style) that returns the actual array of quotes. The template is applied with: var jEl = $("#stockTemplate").tmpl(quotes); which selects the template script tag and uses the .tmpl() function to apply the data to it. The result is a jQuery matched set of elements that can then be appended to the quote display element in the page. The template is merged against an array in this example. When the result is an array the template is automatically applied to each each array item. If you pass a single data item – like say a stock quote – the template works exactly the same way but is applied only once. Templates also have access to a $data item which provides the current data item and information about the tempalte that is currently executing. This makes it possible to keep context within the context of the template itself and also to pass context from a parent template to a child template which is very powerful. Templates can be evaluated by using the template selector and calling the .tmpl() function on the jQuery matched set as shown above or you can use the static $.tmpl() function to provide a template as a string. This allows you to dynamically create templates in code or – more likely – to load templates from the server via AJAX calls. In short there are options The above shows off some of the basics, but there’s much for functionality available in the template engine. Check the documentation link for more information and links to additional examples. The plug-in download also comes with a number of examples that demonstrate functionality. jQuery templates will become a native component in jQuery Core 1.5, so it’s definitely worthwhile checking out the engine today and get familiar with this interface. As much as I’m stoked about templating becoming part of the jQuery core because it’s such an integral part of many applications, there are also a couple shortcomings in the current incarnation: Lack of Error Handling Currently if you embed an expression that is invalid it’s simply not rendered. There’s no error rendered into the template nor do the various  template functions throw errors which leaves finding of bugs as a runtime exercise. I would like some mechanism – optional if possible – to be able to get error info of what is failing in a template when it’s rendered. No String Output Templates are always rendered into a jQuery matched set and there’s no way that I can see to directly render to a string. String output can be useful for debugging as well as opening up templating for creating non-HTML string output. Limited JavaScript Access Unlike John Resig’s original MicroTemplating Engine which was entirely based on JavaScript code generation these templates are limited to a few structured commands that can ‘execute’. There’s no code execution inside of script code which means you’re limited to calling expressions available in global objects or the data item passed in. This may or may not be a big deal depending on the complexity of your template logic. Error handling has been discussed quite a bit and it’s likely there will be some solution to that particualar issue by the time jQuery templates ship. The others are relatively minor issues but something to think about anyway. jQuery Data Link http://api.jquery.com/category/plugins/data-link/ jQuery Data Link provides the ability to do two-way data binding between input controls and an underlying object’s properties. The typical scenario is linking a textbox to a property of an object and have the object updated when the text in the textbox is changed and have the textbox change when the value in the object or the entire object changes. The plug-in also supports converter functions that can be applied to provide the conversion logic from string to some other value typically necessary for mapping things like textbox string input to say a number property and potentially applying additional formatting and calculations. In theory this sounds great, however in reality this plug-in has some serious usability issues. Using the plug-in you can do things like the following to bind data: person = { firstName: "rick", lastName: "strahl"}; $(document).ready( function() { // provide for two-way linking of inputs $("form").link(person); // bind to non-input elements explicitly $("#objFirst").link(person, { firstName: { name: "objFirst", convertBack: function (value, source, target) { $(target).text(value); } } }); $("#objLast").link(person, { lastName: { name: "objLast", convertBack: function (value, source, target) { $(target).text(value); } } }); }); This code hooks up two-way linking between a couple of textboxes on the page and the person object. The first line in the .ready() handler provides mapping of object to form field with the same field names as properties on the object. Note that .link() does NOT bind items into the textboxes when you call .link() – changes are mapped only when values change and you move out of the field. Strike one. The two following commands allow manual binding of values to specific DOM elements which is effectively a one-way bind. You specify the object and a then an explicit mapping where name is an ID in the document. The converter is required to explicitly assign the value to the element. Strike two. You can also detect changes to the underlying object and cause updates to the input elements bound. Unfortunately the syntax to do this is not very natural as you have to rely on the jQuery data object. To update an object’s properties and get change notification looks like this: function updateFirstName() { $(person).data("firstName", person.firstName + " (code updated)"); } This works fine in causing any linked fields to be updated. In the bindings above both the firstName input field and objFirst DOM element gets updated. But the syntax requires you to use a jQuery .data() call for each property change to ensure that the changes are tracked properly. Really? Sure you’re binding through multiple layers of abstraction now but how is that better than just manually assigning values? The code savings (if any) are going to be minimal. As much as I would like to have a WPF/Silverlight/Observable-like binding mechanism in client script, this plug-in doesn’t help much towards that goal in its current incarnation. While you can bind values, the ‘binder’ is too limited to be really useful. If initial values can’t be assigned from the mappings you’re going to end up duplicating work loading the data using some other mechanism. There’s no easy way to re-bind data with a different object altogether since updates trigger only through the .data members. Finally, any non-input elements have to be bound via code that’s fairly verbose and frankly may be more voluminous than what you might write by hand for manual binding and unbinding. Two way binding can be very useful but it has to be easy and most importantly natural. If it’s more work to hook up a binding than writing a couple of lines to do binding/unbinding this sort of thing helps very little in most scenarios. In talking to some of the developers the feature set for Data Link is not complete and they are still soliciting input for features and functionality. If you have ideas on how you want this feature to be more useful get involved and post your recommendations. As it stands, it looks to me like this component needs a lot of love to become useful. For this component to really provide value, bindings need to be able to be refreshed easily and work at the object level, not just the property level. It seems to me we would be much better served by a model binder object that can perform these binding/unbinding tasks in bulk rather than a tool where each link has to be mapped first. I also find the choice of creating a jQuery plug-in questionable – it seems a standalone object – albeit one that relies on the jQuery library – would provide a more intuitive interface than the current forcing of options onto a plug-in style interface. Out of the three Microsoft created components this is by far the least useful and least polished implementation at this point. jQuery Globalization http://github.com/jquery/jquery-global Globalization in JavaScript applications often gets short shrift and part of the reason for this is that natively in JavaScript there’s little support for formatting and parsing of numbers and dates. There are a number of JavaScript libraries out there that provide some support for globalization, but most are limited to a particular portion of globalization. As .NET developers we’re fairly spoiled by the richness of APIs provided in the framework and when dealing with client development one really notices the lack of these features. While you may not necessarily need to localize your application the globalization plug-in also helps with some basic tasks for non-localized applications: Dealing with formatting and parsing of dates and time values. Dates in particular are problematic in JavaScript as there are no formatters whatsoever except the .toString() method which outputs a verbose and next to useless long string. With the globalization plug-in you get a good chunk of the formatting and parsing functionality that the .NET framework provides on the server. You can write code like the following for example to format numbers and dates: var date = new Date(); var output = $.format(date, "MMM. dd, yy") + "\r\n" + $.format(date, "d") + "\r\n" + // 10/25/2010 $.format(1222.32213, "N2") + "\r\n" + $.format(1222.33, "c") + "\r\n"; alert(output); This becomes even more useful if you combine it with templates which can also include any JavaScript expressions. Assuming the globalization plug-in is loaded you can create template expressions that use the $.format function. Here’s the template I used earlier for the stock quote again with a couple of formats applied: <script id="stockTemplate" type="text/x-jquery-tmpl"> <div id="divStockQuote" class="errordisplay" style="width: 500px;"> <div class="label">Company:</div><div><b>${Company}(${Symbol})</b></div> <div class="label">Last Price:</div> <div>${$.format(LastPrice,"N2")}</div> <div class="label">Net Change:</div><div> {{if NetChange > 0}} <b style="color:green" >${NetChange}</b> {{else}} <b style="color:red" >${NetChange}</b> {{/if}} </div> <div class="label">Last Update:</div> <div>${$.format(LastQuoteTime,"MMM dd, yyyy")}</div> </div> </script> There are also parsing methods that can parse dates and numbers from strings into numbers easily: alert($.parseDate("25.10.2010")); alert($.parseInt("12.222")); // de-DE uses . for thousands separators As you can see culture specific options are taken into account when parsing. The globalization plugin provides rich support for a variety of locales: Get a list of all available cultures Query cultures for culture items (like currency symbol, separators etc.) Localized string names for all calendar related items (days of week, months) Generated off of .NET’s supported locales In short you get much of the same functionality that you already might be using in .NET on the server side. The plugin includes a huge number of locales and an Globalization.all.min.js file that contains the text defaults for each of these locales as well as small locale specific script files that define each of the locale specific settings. It’s highly recommended that you NOT use the huge globalization file that includes all locales, but rather add script references to only those languages you explicitly care about. Overall this plug-in is a welcome helper. Even if you use it with a single locale (like en-US) and do no other localization, you’ll gain solid support for number and date formatting which is a vital feature of many applications. Changes for Microsoft It’s good to see Microsoft coming out of its shell and away from the ‘not-built-here’ mentality that has been so pervasive in the past. It’s especially good to see it applied to jQuery – a technology that has stood in drastic contrast to Microsoft’s own internal efforts in terms of design, usage model and… popularity. It’s great to see that Microsoft is paying attention to what customers prefer to use and supporting the customer sentiment – even if it meant drastically changing course of policy and moving into a more open and sharing environment in the process. The additional jQuery support that has been introduced in the last two years certainly has made lives easier for many developers on the ASP.NET platform. It’s also nice to see Microsoft submitting proposals through the standard jQuery process of plug-ins and getting accepted for various very useful projects. Certainly the jQuery Templates plug-in is going to be very useful to many especially since it will be baked into the jQuery core in jQuery 1.5. I hope we see more of this type of involvement from Microsoft in the future. Kudos!© Rick Strahl, West Wind Technologies, 2005-2010Posted in jQuery  ASP.NET  

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  • SimpleMembership, Membership Providers, Universal Providers and the new ASP.NET 4.5 Web Forms and ASP.NET MVC 4 templates

    - by Jon Galloway
    The ASP.NET MVC 4 Internet template adds some new, very useful features which are built on top of SimpleMembership. These changes add some great features, like a much simpler and extensible membership API and support for OAuth. However, the new account management features require SimpleMembership and won't work against existing ASP.NET Membership Providers. I'll start with a summary of top things you need to know, then dig into a lot more detail. Summary: SimpleMembership has been designed as a replacement for traditional the previous ASP.NET Role and Membership provider system SimpleMembership solves common problems people ran into with the Membership provider system and was designed for modern user / membership / storage needs SimpleMembership integrates with the previous membership system, but you can't use a MembershipProvider with SimpleMembership The new ASP.NET MVC 4 Internet application template AccountController requires SimpleMembership and is not compatible with previous MembershipProviders You can continue to use existing ASP.NET Role and Membership providers in ASP.NET 4.5 and ASP.NET MVC 4 - just not with the ASP.NET MVC 4 AccountController The existing ASP.NET Role and Membership provider system remains supported as is part of the ASP.NET core ASP.NET 4.5 Web Forms does not use SimpleMembership; it implements OAuth on top of ASP.NET Membership The ASP.NET Web Site Administration Tool (WSAT) is not compatible with SimpleMembership The following is the result of a few conversations with Erik Porter (PM for ASP.NET MVC) to make sure I had some the overall details straight, combined with a lot of time digging around in ILSpy and Visual Studio's assembly browsing tools. SimpleMembership: The future of membership for ASP.NET The ASP.NET Membership system was introduces with ASP.NET 2.0 back in 2005. It was designed to solve common site membership requirements at the time, which generally involved username / password based registration and profile storage in SQL Server. It was designed with a few extensibility mechanisms - notably a provider system (which allowed you override some specifics like backing storage) and the ability to store additional profile information (although the additional  profile information was packed into a single column which usually required access through the API). While it's sometimes frustrating to work with, it's held up for seven years - probably since it handles the main use case (username / password based membership in a SQL Server database) smoothly and can be adapted to most other needs (again, often frustrating, but it can work). The ASP.NET Web Pages and WebMatrix efforts allowed the team an opportunity to take a new look at a lot of things - e.g. the Razor syntax started with ASP.NET Web Pages, not ASP.NET MVC. The ASP.NET Web Pages team designed SimpleMembership to (wait for it) simplify the task of dealing with membership. As Matthew Osborn said in his post Using SimpleMembership With ASP.NET WebPages: With the introduction of ASP.NET WebPages and the WebMatrix stack our team has really be focusing on making things simpler for the developer. Based on a lot of customer feedback one of the areas that we wanted to improve was the built in security in ASP.NET. So with this release we took that time to create a new built in (and default for ASP.NET WebPages) security provider. I say provider because the new stuff is still built on the existing ASP.NET framework. So what do we call this new hotness that we have created? Well, none other than SimpleMembership. SimpleMembership is an umbrella term for both SimpleMembership and SimpleRoles. Part of simplifying membership involved fixing some common problems with ASP.NET Membership. Problems with ASP.NET Membership ASP.NET Membership was very obviously designed around a set of assumptions: Users and user information would most likely be stored in a full SQL Server database or in Active Directory User and profile information would be optimized around a set of common attributes (UserName, Password, IsApproved, CreationDate, Comment, Role membership...) and other user profile information would be accessed through a profile provider Some problems fall out of these assumptions. Requires Full SQL Server for default cases The default, and most fully featured providers ASP.NET Membership providers (SQL Membership Provider, SQL Role Provider, SQL Profile Provider) require full SQL Server. They depend on stored procedure support, and they rely on SQL Server cache dependencies, they depend on agents for clean up and maintenance. So the main SQL Server based providers don't work well on SQL Server CE, won't work out of the box on SQL Azure, etc. Note: Cory Fowler recently let me know about these Updated ASP.net scripts for use with Microsoft SQL Azure which do support membership, personalization, profile, and roles. But the fact that we need a support page with a set of separate SQL scripts underscores the underlying problem. Aha, you say! Jon's forgetting the Universal Providers, a.k.a. System.Web.Providers! Hold on a bit, we'll get to those... Custom Membership Providers have to work with a SQL-Server-centric API If you want to work with another database or other membership storage system, you need to to inherit from the provider base classes and override a bunch of methods which are tightly focused on storing a MembershipUser in a relational database. It can be done (and you can often find pretty good ones that have already been written), but it's a good amount of work and often leaves you with ugly code that has a bunch of System.NotImplementedException fun since there are a lot of methods that just don't apply. Designed around a specific view of users, roles and profiles The existing providers are focused on traditional membership - a user has a username and a password, some specific roles on the site (e.g. administrator, premium user), and may have some additional "nice to have" optional information that can be accessed via an API in your application. This doesn't fit well with some modern usage patterns: In OAuth and OpenID, the user doesn't have a password Often these kinds of scenarios map better to user claims or rights instead of monolithic user roles For many sites, profile or other non-traditional information is very important and needs to come from somewhere other than an API call that maps to a database blob What would work a lot better here is a system in which you were able to define your users, rights, and other attributes however you wanted and the membership system worked with your model - not the other way around. Requires specific schema, overflow in blob columns I've already mentioned this a few times, but it bears calling out separately - ASP.NET Membership focuses on SQL Server storage, and that storage is based on a very specific database schema. SimpleMembership as a better membership system As you might have guessed, SimpleMembership was designed to address the above problems. Works with your Schema As Matthew Osborn explains in his Using SimpleMembership With ASP.NET WebPages post, SimpleMembership is designed to integrate with your database schema: All SimpleMembership requires is that there are two columns on your users table so that we can hook up to it – an “ID” column and a “username” column. The important part here is that they can be named whatever you want. For instance username doesn't have to be an alias it could be an email column you just have to tell SimpleMembership to treat that as the “username” used to log in. Matthew's example shows using a very simple user table named Users (it could be named anything) with a UserID and Username column, then a bunch of other columns he wanted in his app. Then we point SimpleMemberhip at that table with a one-liner: WebSecurity.InitializeDatabaseFile("SecurityDemo.sdf", "Users", "UserID", "Username", true); No other tables are needed, the table can be named anything we want, and can have pretty much any schema we want as long as we've got an ID and something that we can map to a username. Broaden database support to the whole SQL Server family While SimpleMembership is not database agnostic, it works across the SQL Server family. It continues to support full SQL Server, but it also works with SQL Azure, SQL Server CE, SQL Server Express, and LocalDB. Everything's implemented as SQL calls rather than requiring stored procedures, views, agents, and change notifications. Note that SimpleMembership still requires some flavor of SQL Server - it won't work with MySQL, NoSQL databases, etc. You can take a look at the code in WebMatrix.WebData.dll using a tool like ILSpy if you'd like to see why - there places where SQL Server specific SQL statements are being executed, especially when creating and initializing tables. It seems like you might be able to work with another database if you created the tables separately, but I haven't tried it and it's not supported at this point. Note: I'm thinking it would be possible for SimpleMembership (or something compatible) to run Entity Framework so it would work with any database EF supports. That seems useful to me - thoughts? Note: SimpleMembership has the same database support - anything in the SQL Server family - that Universal Providers brings to the ASP.NET Membership system. Easy to with Entity Framework Code First The problem with with ASP.NET Membership's system for storing additional account information is that it's the gate keeper. That means you're stuck with its schema and accessing profile information through its API. SimpleMembership flips that around by allowing you to use any table as a user store. That means you're in control of the user profile information, and you can access it however you'd like - it's just data. Let's look at a practical based on the AccountModel.cs class in an ASP.NET MVC 4 Internet project. Here I'm adding a Birthday property to the UserProfile class. [Table("UserProfile")] public class UserProfile { [Key] [DatabaseGeneratedAttribute(DatabaseGeneratedOption.Identity)] public int UserId { get; set; } public string UserName { get; set; } public DateTime Birthday { get; set; } } Now if I want to access that information, I can just grab the account by username and read the value. var context = new UsersContext(); var username = User.Identity.Name; var user = context.UserProfiles.SingleOrDefault(u => u.UserName == username); var birthday = user.Birthday; So instead of thinking of SimpleMembership as a big membership API, think of it as something that handles membership based on your user database. In SimpleMembership, everything's keyed off a user row in a table you define rather than a bunch of entries in membership tables that were out of your control. How SimpleMembership integrates with ASP.NET Membership Okay, enough sales pitch (and hopefully background) on why things have changed. How does this affect you? Let's start with a diagram to show the relationship (note: I've simplified by removing a few classes to show the important relationships): So SimpleMembershipProvider is an implementaiton of an ExtendedMembershipProvider, which inherits from MembershipProvider and adds some other account / OAuth related things. Here's what ExtendedMembershipProvider adds to MembershipProvider: The important thing to take away here is that a SimpleMembershipProvider is a MembershipProvider, but a MembershipProvider is not a SimpleMembershipProvider. This distinction is important in practice: you cannot use an existing MembershipProvider (including the Universal Providers found in System.Web.Providers) with an API that requires a SimpleMembershipProvider, including any of the calls in WebMatrix.WebData.WebSecurity or Microsoft.Web.WebPages.OAuth.OAuthWebSecurity. However, that's as far as it goes. Membership Providers still work if you're accessing them through the standard Membership API, and all of the core stuff  - including the AuthorizeAttribute, role enforcement, etc. - will work just fine and without any change. Let's look at how that affects you in terms of the new templates. Membership in the ASP.NET MVC 4 project templates ASP.NET MVC 4 offers six Project Templates: Empty - Really empty, just the assemblies, folder structure and a tiny bit of basic configuration. Basic - Like Empty, but with a bit of UI preconfigured (css / images / bundling). Internet - This has both a Home and Account controller and associated views. The Account Controller supports registration and login via either local accounts and via OAuth / OpenID providers. Intranet - Like the Internet template, but it's preconfigured for Windows Authentication. Mobile - This is preconfigured using jQuery Mobile and is intended for mobile-only sites. Web API - This is preconfigured for a service backend built on ASP.NET Web API. Out of these templates, only one (the Internet template) uses SimpleMembership. ASP.NET MVC 4 Basic template The Basic template has configuration in place to use ASP.NET Membership with the Universal Providers. You can see that configuration in the ASP.NET MVC 4 Basic template's web.config: <profile defaultProvider="DefaultProfileProvider"> <providers> <add name="DefaultProfileProvider" type="System.Web.Providers.DefaultProfileProvider, System.Web.Providers, Version=1.0.0.0, Culture=neutral, PublicKeyToken=31bf3856ad364e35" connectionStringName="DefaultConnection" applicationName="/" /> </providers> </profile> <membership defaultProvider="DefaultMembershipProvider"> <providers> <add name="DefaultMembershipProvider" type="System.Web.Providers.DefaultMembershipProvider, System.Web.Providers, Version=1.0.0.0, Culture=neutral, PublicKeyToken=31bf3856ad364e35" connectionStringName="DefaultConnection" enablePasswordRetrieval="false" enablePasswordReset="true" requiresQuestionAndAnswer="false" requiresUniqueEmail="false" maxInvalidPasswordAttempts="5" minRequiredPasswordLength="6" minRequiredNonalphanumericCharacters="0" passwordAttemptWindow="10" applicationName="/" /> </providers> </membership> <roleManager defaultProvider="DefaultRoleProvider"> <providers> <add name="DefaultRoleProvider" type="System.Web.Providers.DefaultRoleProvider, System.Web.Providers, Version=1.0.0.0, Culture=neutral, PublicKeyToken=31bf3856ad364e35" connectionStringName="DefaultConnection" applicationName="/" /> </providers> </roleManager> <sessionState mode="InProc" customProvider="DefaultSessionProvider"> <providers> <add name="DefaultSessionProvider" type="System.Web.Providers.DefaultSessionStateProvider, System.Web.Providers, Version=1.0.0.0, Culture=neutral, PublicKeyToken=31bf3856ad364e35" connectionStringName="DefaultConnection" /> </providers> </sessionState> This means that it's business as usual for the Basic template as far as ASP.NET Membership works. ASP.NET MVC 4 Internet template The Internet template has a few things set up to bootstrap SimpleMembership: \Models\AccountModels.cs defines a basic user account and includes data annotations to define keys and such \Filters\InitializeSimpleMembershipAttribute.cs creates the membership database using the above model, then calls WebSecurity.InitializeDatabaseConnection which verifies that the underlying tables are in place and marks initialization as complete (for the application's lifetime) \Controllers\AccountController.cs makes heavy use of OAuthWebSecurity (for OAuth account registration / login / management) and WebSecurity. WebSecurity provides account management services for ASP.NET MVC (and Web Pages) WebSecurity can work with any ExtendedMembershipProvider. There's one in the box (SimpleMembershipProvider) but you can write your own. Since a standard MembershipProvider is not an ExtendedMembershipProvider, WebSecurity will throw exceptions if the default membership provider is a MembershipProvider rather than an ExtendedMembershipProvider. Practical example: Create a new ASP.NET MVC 4 application using the Internet application template Install the Microsoft ASP.NET Universal Providers for LocalDB NuGet package Run the application, click on Register, add a username and password, and click submit You'll get the following execption in AccountController.cs::Register: To call this method, the "Membership.Provider" property must be an instance of "ExtendedMembershipProvider". This occurs because the ASP.NET Universal Providers packages include a web.config transform that will update your web.config to add the Universal Provider configuration I showed in the Basic template example above. When WebSecurity tries to use the configured ASP.NET Membership Provider, it checks if it can be cast to an ExtendedMembershipProvider before doing anything else. So, what do you do? Options: If you want to use the new AccountController, you'll either need to use the SimpleMembershipProvider or another valid ExtendedMembershipProvider. This is pretty straightforward. If you want to use an existing ASP.NET Membership Provider in ASP.NET MVC 4, you can't use the new AccountController. You can do a few things: Replace  the AccountController.cs and AccountModels.cs in an ASP.NET MVC 4 Internet project with one from an ASP.NET MVC 3 application (you of course won't have OAuth support). Then, if you want, you can go through and remove other things that were built around SimpleMembership - the OAuth partial view, the NuGet packages (e.g. the DotNetOpenAuthAuth package, etc.) Use an ASP.NET MVC 4 Internet application template and add in a Universal Providers NuGet package. Then copy in the AccountController and AccountModel classes. Create an ASP.NET MVC 3 project and upgrade it to ASP.NET MVC 4 using the steps shown in the ASP.NET MVC 4 release notes. None of these are particularly elegant or simple. Maybe we (or just me?) can do something to make this simpler - perhaps a NuGet package. However, this should be an edge case - hopefully the cases where you'd need to create a new ASP.NET but use legacy ASP.NET Membership Providers should be pretty rare. Please let me (or, preferably the team) know if that's an incorrect assumption. Membership in the ASP.NET 4.5 project template ASP.NET 4.5 Web Forms took a different approach which builds off ASP.NET Membership. Instead of using the WebMatrix security assemblies, Web Forms uses Microsoft.AspNet.Membership.OpenAuth assembly. I'm no expert on this, but from a bit of time in ILSpy and Visual Studio's (very pretty) dependency graphs, this uses a Membership Adapter to save OAuth data into an EF managed database while still running on top of ASP.NET Membership. Note: There may be a way to use this in ASP.NET MVC 4, although it would probably take some plumbing work to hook it up. How does this fit in with Universal Providers (System.Web.Providers)? Just to summarize: Universal Providers are intended for cases where you have an existing ASP.NET Membership Provider and you want to use it with another SQL Server database backend (other than SQL Server). It doesn't require agents to handle expired session cleanup and other background tasks, it piggybacks these tasks on other calls. Universal Providers are not really, strictly speaking, universal - at least to my way of thinking. They only work with databases in the SQL Server family. Universal Providers do not work with Simple Membership. The Universal Providers packages include some web config transforms which you would normally want when you're using them. What about the Web Site Administration Tool? Visual Studio includes tooling to launch the Web Site Administration Tool (WSAT) to configure users and roles in your application. WSAT is built to work with ASP.NET Membership, and is not compatible with Simple Membership. There are two main options there: Use the WebSecurity and OAuthWebSecurity API to manage the users and roles Create a web admin using the above APIs Since SimpleMembership runs on top of your database, you can update your users as you would any other data - via EF or even in direct database edits (in development, of course)

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