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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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• 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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• Nagios plugin script not working as expected

  - by Linker3000

  I have modified an off-the-shelf Nagios plugin perl script to (in theory) return a one or zero according to the existence, or not, of a file on a remote linux server. The script runs a remote ssh session and logs in as the nagios user. The remote linux servers have private keys setup for that user, and on the bash command line the script works as expected, but when run as a plugin it always returns '1' (true) even if the file does not exist. Some help with the logic or a comment on why things are not working as expected within Nagios would be appreciated. I'd prefer to use this ssh login method rather than having to install nrpe on all the linux servers. To run from a command line (assuming remote server has a user called nagios with a valid private key): ./check_reboot_required -e ssh -H remote-servers-ip-addr -p 'filename-to-check' -v Ta. #! /usr/bin/perl -w # # # License Information: # This program is free software; you can redistribute it and/or modify # it under the terms of the GNU General Public License as published by # the Free Software Foundation; either version 2 of the License, or # (at your option) any later version. # # This program is distributed in the hope that it will be useful, # but WITHOUT ANY WARRANTY; without even the implied warranty of # MERCHANTABILITY or FITNESS FOR A PARTICULAR PURPOSE. See the # GNU General Public License for more details. # # You should have received a copy of the GNU General Public License # along with this program; if not, write to the Free Software # Foundation, Inc., 675 Mass Ave, Cambridge, MA 02139, USA. # ############################################################################ use POSIX; use strict; use Getopt::Long; use lib "/usr/lib/nagios/plugins" ; use vars qw($host $opt_V $opt_h $opt_v $verbose $PROGNAME $pattern $opt_p $mmin $opt_e $opt_t $opt_H $status $state $msg $msg_q $MAILQ $SHELL $device $used $avail $percent $fs $blocks $CMD $RMTOS); use utils qw(%ERRORS &print_revision &support &usage ); sub print_help (); sub print_usage (); sub process_arguments (); $ENV{'PATH'}=''; $ENV{'BASH_ENV'}=''; $ENV{'ENV'}=''; $PROGNAME = "check_reboot_required"; Getopt::Long::Configure('bundling'); $status = process_arguments(); if ($status){ print "ERROR: processing arguments\n"; exit $ERRORS{'UNKNOWN'}; } $SIG{'ALRM'} = sub { print ("ERROR: timed out waiting for $CMD on $host\n"); exit $ERRORS{'WARNING'}; }; $host = $opt_H; $pattern = $opt_p; print "Pattern >" . $pattern . "< " if $verbose; alarm($opt_t); #$CMD = "/usr/bin/find " . $pattern . " -type f 2>/dev/null| /usr/bin/wc -l"; $CMD = "[ -f " . $pattern . " ] && echo 1 || echo 0"; alarm($opt_t); ## get cmd output from remote system if (! open (OUTPUT, "$SHELL $host $CMD|" ) ) { print "ERROR: could not open $CMD on $host\n"; exit $ERRORS{'UNKNOWN'}; } my $perfdata = ""; my $state = "3"; my $msg = "Indeterminate result"; # only first line is relevant in this iteration. while (<OUTPUT>) { my $result = chomp($_); $msg = $result; print "Shell returned >" . $result . "< length is " . length($result) . " " if $verbose; if ( $result == 1 ) { $msg = "Reboot required (NB: Result still not accurate)" . $result ; $state = $ERRORS{'WARNING'}; last; } elsif ( $result == 0 ) { $msg = "No reboot required (NB: Result still not accurate) " . $result ; $state = $ERRORS{'OK'}; last; } else { $msg = "Output received, but it was neither a 1 nor a 0" ; last; } } close (OUTPUT); print "$msg | $perfdata\n"; exit $state; ##################################### #### subs sub process_arguments(){ GetOptions ("V" => \$opt_V, "version" => \$opt_V, "v" => \$opt_v, "verbose" => \$opt_v, "h" => \$opt_h, "help" => \$opt_h, "e=s" => \$opt_e, "shell=s" => \$opt_e, "p=s" => \$opt_p, "pattern=s" => \$opt_p, "t=i" => \$opt_t, "timeout=i" => \$opt_t, "H=s" => \$opt_H, "hostname=s" => \$opt_H ); if ($opt_V) { print_revision($PROGNAME,'$Revision: 1.0 $ '); exit $ERRORS{'OK'}; } if ($opt_h) { print_help(); exit $ERRORS{'OK'}; } if (defined $opt_v ){ $verbose = $opt_v; } if (defined $opt_e ){ if ( $opt_e eq "ssh" ) { if (-x "/usr/local/bin/ssh") { $SHELL = "/usr/local/bin/ssh"; } elsif ( -x "/usr/bin/ssh" ) { $SHELL = "/usr/bin/ssh"; } else { print_usage(); exit $ERRORS{'UNKNOWN'}; } } elsif ( $opt_e eq "rsh" ) { $SHELL = "/usr/bin/rsh"; } else { print_usage(); exit $ERRORS{'UNKNOWN'}; } } else { print_usage(); exit $ERRORS{'UNKNOWN'}; } unless (defined $opt_t) { $opt_t = $utils::TIMEOUT ; # default timeout } unless (defined $opt_H) { print_usage(); exit $ERRORS{'UNKNOWN'}; } return $ERRORS{'OK'}; } sub print_usage () { print "Usage: $PROGNAME -e <shell> -H <hostname> -p <directory/file pattern> [-t <timeout>] [-v verbose]\n"; } sub print_help () { print_revision($PROGNAME,'$Revision: 0.1 $'); print "\n"; print_usage(); print "\n"; print " Checks for the presence of a 'reboot-required' file on a remote host via SSH or RSH\n"; print "-e (--shell) = ssh or rsh (required)\n"; print "-H (--hostname) = remote server name (required)"; print "-p (--pattern) = File pattern for find command (default = /var/run/reboot-required)\n"; print "-t (--timeout) = Plugin timeout in seconds (default = $utils::TIMEOUT)\n"; print "-h (--help)\n"; print "-V (--version)\n"; print "-v (--verbose) = debugging output\n"; print "\n\n"; support(); }

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• Linux server is only using 60% of memory, then swapping

  - by Kamil Kisiel

  I've got a Linux server that's running our bacula backup system. The machine is grinding like mad because it's going heavy in to swap. The problem is, it's only using 60% of its physical memory! Here's the output from free -m: free -m total used free shared buffers cached Mem: 3949 2356 1593 0 0 1 -/+ buffers/cache: 2354 1595 Swap: 7629 1804 5824 and some sample output from vmstat 1: procs -----------memory---------- ---swap-- -----io---- -system-- -----cpu------ r b swpd free buff cache si so bi bo in cs us sy id wa st 0 2 1843536 1634512 0 4188 54 13 2524 666 2 1 1 1 89 9 0 1 11 1845916 1640724 0 388 2700 4816 221880 4879 14409 170721 4 3 63 30 0 0 9 1846096 1643952 0 0 4956 756 174832 804 12357 159306 3 4 63 30 0 0 11 1846104 1643532 0 0 4916 540 174320 580 10609 139960 3 4 64 29 0 0 4 1846084 1640272 0 2336 4080 524 140408 548 9331 118287 3 4 63 30 0 0 8 1846104 1642096 0 1488 2940 432 102516 457 7023 82230 2 4 65 29 0 0 5 1846104 1642268 0 1276 3704 452 126520 452 9494 119612 3 5 65 27 0 3 12 1846104 1641528 0 328 6092 608 187776 636 8269 113059 4 3 64 29 0 2 2 1846084 1640960 0 724 5948 0 111480 0 7751 116370 4 4 63 29 0 0 4 1846100 1641484 0 404 4144 1476 125760 1500 10668 105358 2 3 71 25 0 0 13 1846104 1641932 0 0 5872 828 153808 840 10518 128447 3 4 70 22 0 0 8 1846096 1639172 0 3164 3556 556 74884 580 5082 65362 2 2 73 23 0 1 4 1846080 1638676 0 396 4512 28 50928 44 2672 38277 2 2 80 16 0 0 3 1846080 1628808 0 7132 2636 0 28004 8 1358 14090 0 1 78 20 0 0 2 1844728 1618552 0 11140 7680 0 12740 8 763 2245 0 0 82 18 0 0 2 1837764 1532056 0 101504 2952 0 95644 24 802 3817 0 1 87 12 0 0 11 1842092 1633324 0 4416 1748 10900 143144 11024 6279 134442 3 3 70 24 0 2 6 1846104 1642756 0 0 4768 468 78752 468 4672 60141 2 2 76 20 0 1 12 1846104 1640792 0 236 4752 440 140712 464 7614 99593 3 5 58 34 0 0 3 1846084 1630368 0 6316 5104 0 20336 0 1703 22424 1 1 72 26 0 2 17 1846104 1638332 0 3168 4080 1720 211960 1744 11977 155886 3 4 65 28 0 1 10 1846104 1640800 0 132 4488 556 126016 584 8016 106368 3 4 63 29 0 0 14 1846104 1639740 0 2248 3436 428 114188 452 7030 92418 3 3 59 35 0 1 6 1846096 1639504 0 1932 5500 436 141412 460 8261 112210 4 4 63 29 0 0 10 1846104 1640164 0 3052 4028 448 147684 472 7366 109554 4 4 61 30 0 0 10 1846100 1641040 0 2332 4952 632 147452 664 8767 118384 3 4 63 30 0 4 8 1846084 1641092 0 664 4948 276 152264 292 6448 98813 5 5 62 28 0 Furthermore, the output of top sorted by CPU time seems to support the theory that swap is what's bogging down the system: top - 09:05:32 up 37 days, 23:24, 1 user, load average: 9.75, 8.24, 7.12 Tasks: 173 total, 1 running, 172 sleeping, 0 stopped, 0 zombie Cpu(s): 1.6%us, 1.4%sy, 0.0%ni, 76.1%id, 20.6%wa, 0.1%hi, 0.2%si, 0.0%st Mem: 4044632k total, 2405628k used, 1639004k free, 0k buffers Swap: 7812492k total, 1851852k used, 5960640k free, 436k cached PID USER PR NI VIRT RES SHR S %CPU %MEM TIME+ TIME COMMAND 4174 root 17 0 63156 176 56 S 8 0.0 2138:52 35,38 bacula-fd 4185 root 17 0 63352 284 104 S 6 0.0 1709:25 28,29 bacula-sd 240 root 15 0 0 0 0 D 3 0.0 831:55.19 831:55 kswapd0 2852 root 10 -5 0 0 0 S 1 0.0 126:35.59 126:35 xfsbufd 2849 root 10 -5 0 0 0 S 0 0.0 119:50.94 119:50 xfsbufd 1364 root 10 -5 0 0 0 S 0 0.0 117:05.39 117:05 xfsbufd 21 root 10 -5 0 0 0 S 1 0.0 48:03.44 48:03 events/3 6940 postgres 16 0 43596 8 8 S 0 0.0 46:50.35 46:50 postmaster 1342 root 10 -5 0 0 0 S 0 0.0 23:14.34 23:14 xfsdatad/4 5415 root 17 0 1770m 108 48 S 0 0.0 15:03.74 15:03 bacula-dir 23 root 10 -5 0 0 0 S 0 0.0 13:09.71 13:09 events/5 5604 root 17 0 1216m 500 200 S 0 0.0 12:38.20 12:38 java 5552 root 16 0 1194m 580 248 S 0 0.0 11:58.00 11:58 java Here's the same sorted by virtual memory image size: top - 09:08:32 up 37 days, 23:27, 1 user, load average: 8.43, 8.26, 7.32 Tasks: 173 total, 1 running, 172 sleeping, 0 stopped, 0 zombie Cpu(s): 3.6%us, 3.4%sy, 0.0%ni, 62.2%id, 30.2%wa, 0.2%hi, 0.3%si, 0.0%st Mem: 4044632k total, 2404212k used, 1640420k free, 0k buffers Swap: 7812492k total, 1852548k used, 5959944k free, 100k cached PID USER PR NI VIRT RES SHR S %CPU %MEM TIME+ TIME COMMAND 5415 root 17 0 1770m 56 44 S 0 0.0 15:03.78 15:03 bacula-dir 5604 root 17 0 1216m 492 200 S 0 0.0 12:38.30 12:38 java 5552 root 16 0 1194m 476 200 S 0 0.0 11:58.20 11:58 java 4598 root 16 0 117m 44 44 S 0 0.0 0:13.37 0:13 eventmond 9614 gdm 16 0 93188 0 0 S 0 0.0 0:00.30 0:00 gdmgreeter 5527 root 17 0 78716 0 0 S 0 0.0 0:00.30 0:00 gdm 4185 root 17 0 63352 284 104 S 20 0.0 1709:52 28,29 bacula-sd 4174 root 17 0 63156 208 88 S 24 0.0 2139:25 35,39 bacula-fd 10849 postgres 18 0 54740 216 108 D 0 0.0 0:31.40 0:31 postmaster 6661 postgres 17 0 49432 0 0 S 0 0.0 0:03.50 0:03 postmaster 5507 root 15 0 47980 0 0 S 0 0.0 0:00.00 0:00 gdm 6940 postgres 16 0 43596 16 16 S 0 0.0 46:51.39 46:51 postmaster 5304 postgres 16 0 40580 132 88 S 0 0.0 6:21.79 6:21 postmaster 5301 postgres 17 0 40448 24 24 S 0 0.0 0:32.17 0:32 postmaster 11280 root 16 0 40288 28 28 S 0 0.0 0:00.11 0:00 sshd 5534 root 17 0 37580 0 0 S 0 0.0 0:56.18 0:56 X 30870 root 30 15 31668 28 28 S 0 0.0 1:13.38 1:13 snmpd 5305 postgres 17 0 30628 16 16 S 0 0.0 0:11.60 0:11 postmaster 27403 postfix 17 0 30248 0 0 S 0 0.0 0:02.76 0:02 qmgr 10815 postfix 15 0 30208 16 16 S 0 0.0 0:00.02 0:00 pickup 5306 postgres 16 0 29760 20 20 S 0 0.0 0:52.89 0:52 postmaster 5302 postgres 17 0 29628 64 32 S 0 0.0 1:00.64 1:00 postmaster I've tried tuning the swappiness kernel parameter to both high and low values, but nothing appears to change the behavior here. I'm at a loss to figure out what's going on. How can I find out what's causing this? Update: The system is a fully 64-bit system, so there should be no question of memory limitations due to 32-bit issues. Update2: As I mentioned in the original question, I've already tried tuning swappiness to all sorts of values, including 0. The result is always the same, with approximately 1.6 GB of memory remaining unused. Update3: Added top output to the above info.

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• Android - HorizontalScrollView within ScrollView Touch Handling

  - by Joel

  Hi, I have a ScrollView that surrounds my entire layout so that the entire screen is scrollable. The first element I have in this ScrollView is a HorizontalScrollView block that has features that can be scrolled through horizontally. I've added an ontouchlistener to the horizontalscrollview to handle touch events and force the view to "snap" to the closest image on the ACTION_UP event. So the effect I'm going for is like the stock android homescreen where you can scroll from one to the other and it snaps to one screen when you lift your finger. This all works great except for one problem: I need to swipe left to right almost perfectly horizontally for an ACTION_UP to ever register. If I swipe vertically in the very least (which I think many people tend to do on their phones when swiping side to side), I will receive an ACTION_CANCEL instead of an ACTION_UP. My theory is that this is because the horizontalscrollview is within a scrollview, and the scrollview is hijacking the vertical touch to allow for vertical scrolling. How can I disable the touch events for the scrollview from just within my horizontal scrollview, but still allow for normal vertical scrolling elsewhere in the scrollview? Here's a sample of my code: public class HomeFeatureLayout extends HorizontalScrollView { private ArrayList<ListItem> items = null; private GestureDetector gestureDetector; View.OnTouchListener gestureListener; private static final int SWIPE_MIN_DISTANCE = 5; private static final int SWIPE_THRESHOLD_VELOCITY = 300; private int activeFeature = 0; public HomeFeatureLayout(Context context, ArrayList<ListItem> items){ super(context); setLayoutParams(new LayoutParams(LayoutParams.FILL_PARENT, LayoutParams.WRAP_CONTENT)); setFadingEdgeLength(0); this.setHorizontalScrollBarEnabled(false); this.setVerticalScrollBarEnabled(false); LinearLayout internalWrapper = new LinearLayout(context); internalWrapper.setLayoutParams(new LayoutParams(LayoutParams.FILL_PARENT, LayoutParams.FILL_PARENT)); internalWrapper.setOrientation(LinearLayout.HORIZONTAL); addView(internalWrapper); this.items = items; for(int i = 0; i< items.size();i++){ LinearLayout featureLayout = (LinearLayout) View.inflate(this.getContext(),R.layout.homefeature,null); TextView header = (TextView) featureLayout.findViewById(R.id.featureheader); ImageView image = (ImageView) featureLayout.findViewById(R.id.featureimage); TextView title = (TextView) featureLayout.findViewById(R.id.featuretitle); title.setTag(items.get(i).GetLinkURL()); TextView date = (TextView) featureLayout.findViewById(R.id.featuredate); header.setText("FEATURED"); Image cachedImage = new Image(this.getContext(), items.get(i).GetImageURL()); image.setImageDrawable(cachedImage.getImage()); title.setText(items.get(i).GetTitle()); date.setText(items.get(i).GetDate()); internalWrapper.addView(featureLayout); } gestureDetector = new GestureDetector(new MyGestureDetector()); setOnTouchListener(new View.OnTouchListener() { @Override public boolean onTouch(View v, MotionEvent event) { if (gestureDetector.onTouchEvent(event)) { return true; } else if(event.getAction() == MotionEvent.ACTION_UP || event.getAction() == MotionEvent.ACTION_CANCEL ){ int scrollX = getScrollX(); int featureWidth = getMeasuredWidth(); activeFeature = ((scrollX + (featureWidth/2))/featureWidth); int scrollTo = activeFeature*featureWidth; smoothScrollTo(scrollTo, 0); return true; } else{ return false; } } }); } class MyGestureDetector extends SimpleOnGestureListener { @Override public boolean onFling(MotionEvent e1, MotionEvent e2, float velocityX, float velocityY) { try { //right to left if(e1.getX() - e2.getX() > SWIPE_MIN_DISTANCE && Math.abs(velocityX) > SWIPE_THRESHOLD_VELOCITY) { activeFeature = (activeFeature < (items.size() - 1))? activeFeature + 1:items.size() -1; smoothScrollTo(activeFeature*getMeasuredWidth(), 0); return true; } //left to right else if (e2.getX() - e1.getX() > SWIPE_MIN_DISTANCE && Math.abs(velocityX) > SWIPE_THRESHOLD_VELOCITY) { activeFeature = (activeFeature > 0)? activeFeature - 1:0; smoothScrollTo(activeFeature*getMeasuredWidth(), 0); return true; } } catch (Exception e) { // nothing } return false; } } }

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• How to reduce iOS AVPlayer start delay

  - by Bernt Habermeier

  Note, for the below question: All assets are local on the device -- no network streaming is taking place. The videos contain audio tracks. I'm working on an iOS application that requires playing video files with minimum delay to start the video clip in question. Unfortunately we do not know what specific video clip is next until we actually need to start it up. Specifically: When one video clip is playing, we will know what the next set of (roughly) 10 video clips are, but we don't know which one exactly, until it comes time to 'immediately' play the next clip. What I've done to look at actual start delays is to call addBoundaryTimeObserverForTimes on the video player, with a time period of one millisecond to see when the video actually started to play, and I take the difference of that time stamp with the first place in the code that indicates which asset to start playing. From what I've seen thus-far, I have found that using the combination of AVAsset loading, and then creating an AVPlayerItem from that once it's ready, and then waiting for AVPlayerStatusReadyToPlay before I call play, tends to take between 1 and 3 seconds to start the clip. I've since switched to what I think is roughly equivalent: calling [AVPlayerItem playerItemWithURL:] and waiting for AVPlayerItemStatusReadyToPlay to play. Roughly same performance. One thing I'm observing is that the first AVPlayer item load is slower than the rest. Seems one idea is to pre-flight the AVPlayer with a short / empty asset before trying to play the first video might be of good general practice. [http://stackoverflow.com/questions/900461/slow-start-for-avaudioplayer-the-first-time-a-sound-is-played] I'd love to get the video start times down as much as possible, and have some ideas of things to experiment with, but would like some guidance from anyone that might be able to help. Update: idea 7, below, as-implemented yields switching times of around 500 ms. This is an improvement, but it it'd be nice to get this even faster. Idea 1: Use N AVPlayers (won't work) Using ~ 10 AVPPlayer objects and start-and-pause all ~ 10 clips, and once we know which one we really need, switch to, and un-pause the correct AVPlayer, and start all over again for the next cycle. I don't think this works, because I've read there is roughly a limit of 4 active AVPlayer's in iOS. There was someone asking about this on StackOverflow here, and found out about the 4 AVPlayer limit: fast-switching-between-videos-using-avfoundation Idea 2: Use AVQueuePlayer (won't work) I don't believe that shoving 10 AVPlayerItems into an AVQueuePlayer would pre-load them all for seamless start. AVQueuePlayer is a queue, and I think it really only makes the next video in the queue ready for immediate playback. I don't know which one out of ~10 videos we do want to play back, until it's time to start that one. ios-avplayer-video-preloading Idea 3: Load, Play, and retain AVPlayerItems in background (not 100% sure yet -- but not looking good) I'm looking at if there is any benefit to load and play the first second of each video clip in the background (suppress video and audio output), and keep a reference to each AVPlayerItem, and when we know which item needs to be played for real, swap that one in, and swap the background AVPlayer with the active one. Rinse and Repeat. The theory would be that recently played AVPlayer/AVPlayerItem's may still hold some prepared resources which would make subsequent playback faster. So far, I have not seen benefits from this, but I might not have the AVPlayerLayer setup correctly for the background. I doubt this will really improve things from what I've seen. Idea 4: Use a different file format -- maybe one that is faster to load? I'm currently using .m4v's (video-MPEG4) H.264 format. I have not played around with other formats, but it may well be that some formats are faster to decode / get ready than others. Possible still using video-MPEG4 but with a different codec, or maybe quicktime? Maybe a lossless video format where decoding / setup is faster? Idea 5: Combination of lossless video format + AVQueuePlayer If there is a video format that is fast to load, but maybe where the file size is insane, one idea might be to pre-prepare the first 10 seconds of each video clip with a version that is boated but faster to load, but back that up with an asset that is encoded in H.264. Use an AVQueuePlayer, and add the first 10 seconds in the uncompressed file format, and follow that up with one that is in H.264 which gets up to 10 seconds of prepare/preload time. So I'd get 'the best' of both worlds: fast start times, but also benefits from a more compact format. Idea 6: Use a non-standard AVPlayer / write my own / use someone else's Given my needs, maybe I can't use AVPlayer, but have to resort to AVAssetReader, and decode the first few seconds (possibly write raw file to disk), and when it comes to playback, make use of the raw format to play it back fast. Seems like a huge project to me, and if I go about it in a naive way, it's unclear / unlikely to even work better. Each decoded and uncompressed video frame is 2.25 MB. Naively speaking -- if we go with ~ 30 fps for the video, I'd end up with ~60 MB/s read-from-disk requirement, which is probably impossible / pushing it. Obviously we'd have to do some level of image compression (perhaps native openGL/es compression formats via PVRTC)... but that's kind crazy. Maybe there is a library out there that I can use? Idea 7: Combine everything into a single movie asset, and seekToTime One idea that might be easier than some of the above, is to combine everything into a single movie, and use seekToTime. The thing is that we'd be jumping all around the place. Essentially random access into the movie. I think this may actually work out okay: avplayer-movie-playing-lag-in-ios5 Which approach do you think would be best? So far, I've not made that much progress in terms of reducing the lag.

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• Getting started with Exchange Web Services 2010

  - by Adam Tuttle

  I've been tasked with writing a SOAP web-service in .Net to be middleware between EWS2010 and an application server that previously used WebDAV to connect to Exchange. (As I understand it, WebDAV is going away with EWS2010, so the application server will no longer be able to connect as it previously did, and it is exponentially harder to connect to EWS without WebDAV. The theory is that doing it in .Net should be easier than anything else... Right?!) My end goal is to be able to get and create/update email, calendar items, contacts, and to-do list items for a specified Exchange account. (Deleting is not currently necessary, but I may build it in for future consideration, if it's easy enough). I was originally given some sample code, which did in fact work, but I quickly realized that it was outdated. The types and classes used appear nowhere in the current documentation. For example, the method used to create a connection to the Exchange server was: ExchangeService svc = new ExchangeService(); svc.Credentials = new WebCredentials(AuthEmailAddress, AuthEmailPassword); svc.AutodiscoverUrl(AutoDiscoverEmailAddress); For what it's worth, this was using an assembly that came with the sample code: Microsoft.Exchange.WebServices.dll ("MEWS"). Before I realized that this wasn't the current standard way to accomplish the connection, and it worked, I tried to build on it and add a method to create calendar items, which I copied from here: static void CreateAppointment(ExchangeServiceBinding esb) { // Create the appointment. CalendarItemType appointment = new CalendarItemType(); ... } Right away, I'm confronted with the difference between ExchangeService and ExchangeServiceBinding ("ESB"); so I started Googling to try and figure out how to get an ESB definition so that the CreateAppointment method will compile. I found this blog post that explains how to generate a proxy class from a WSDL, which I did. Unfortunately, this caused some conflicts where types that were defined in the original Assembly, Microsoft.Exchange.WebServices.dll (that came with the sample code) overlapped with Types in my new EWS.dll assembly (which I compiled from the code generated from the services.wsdl provided by the Exchange server). I excluded the MEWS assembly, which only made things worse. I went from a handful of errors and warnings to 25 errors and 2,510 warnings. All kinds of types and methods were not found. Something is clearly wrong, here. So I went back on the hunt. I found instructions on adding service references and web references (i.e. the extra steps it takes in VS2008), and I think I'm back on the right track. I removed (actually, for now, just excluded) all previous assemblies I had been trying; and I added a service reference for https://my.exchange-server.com/ews/services.wsdl Now I'm down to just 1 error and 1 warning. Warning: The element 'transport' cannot contain child element 'extendedProtectionPolicy' because the parent element's content model is empty. This is in reference to a change that was made to web.config when I added the service reference; and I just found a fix for that here on SO. I've commented that section out as indicated, and it did make the warning go away, so woot for that. The error hasn't been so easy to get around, though: Error: The type or namespace name 'ExchangeService' could not be found (are you missing a using directive or an assembly reference?) This is in reference to the function I was using to create the EWS connection, called by each of the web methods: private ExchangeService getService(String AutoDiscoverEmailAddress, String AuthEmailAddress, String AuthEmailPassword) { ExchangeService svc = new ExchangeService(); svc.Credentials = new WebCredentials(AuthEmailAddress, AuthEmailPassword); svc.AutodiscoverUrl(AutoDiscoverEmailAddress); return svc; } This function worked perfectly with the MEWS assembly from the sample code, but the ExchangeService type is no longer available. (Nor is ExchangeServiceBinding, that was the first thing I checked.) At this point, since I'm not following any directions from the documentation (I couldn't find anywhere in the documentation that said to add a service reference to your Exchange server's services.wsdl -- but that does seem to be the best/farthest I've gotten so far), I feel like I'm flying blind. I know I need to figure out whatever it is that should replace ExchangeService / ExchangeServiceBinding, implement that, and then work through whatever errors crop up as a result of that switch... But I have no idea how to do that, or where to look for how to do it. Googling "ExchangeService" and "ExchangeServiceBinding" only seem to lead back to outdated blog posts and MSDN, neither of which has proven terribly helpful thus far. Help me, Obi-Wan, you're my only hope!

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• NHibernate: Collection was modified; enumeration operation may not execute

  - by Daoming Yang

  Hi All, I'm currently struggling with this "Collection was modified; enumeration operation may not execute" issue. I have searched about this error message, and it's all related to the foreach statement. I do have the some foreach statements, but they are just simply representing the data. I did not using any remove or add inside the foreach statement. NOTE: The error randomly happens (about 4-5 times a day). The application is the MVC website. There are about 5 users operate this applications (about 150 orders a day). Could it be some another users modified the collection, and then occur this error? I have log4net setup and the settings can be found here Make sure that the controller has a parameterless public constructor I do have parameterless public constructor in AdminProductController Does anyone know why this happen and how to resolve this issue? A friend (Oskar) mentioned that "Theory: Maybe the problem is that your configuration and session factory is initialized on the first request after application restart. If a second request comes in before the first request is finished, maybe it will also try to initialize and then triggering this problem somehow." Many thanks. Daoming Here is the error message: System.InvalidOperationException Collection was modified; enumeration operation may not execute. System.InvalidOperationException: An error occurred when trying to create a controller of type 'WebController.Controllers.Admin.AdminProductController'. Make sure that the controller has a parameterless public constructor. --- System.Reflection.TargetInvocationException: Exception has been thrown by the target of an invocation. --- NHibernate.MappingException: Could not configure datastore from input stream DomainModel.Entities.Mappings.OrderProductVariant.hbm.xml --- System.InvalidOperationException: Collection was modified; enumeration operation may not execute. at System.Collections.ArrayList.ArrayListEnumeratorSimple.MoveNext() at System.Xml.Schema.XmlSchemaSet.AddSchemaToSet(XmlSchema schema) at System.Xml.Schema.XmlSchemaSet.Add(String targetNamespace, XmlSchema schema) at System.Xml.Schema.XmlSchemaSet.Add(XmlSchema schema) at NHibernate.Cfg.Configuration.LoadMappingDocument(XmlReader hbmReader, String name) at NHibernate.Cfg.Configuration.AddInputStream(Stream xmlInputStream, String name) --- End of inner exception stack trace --- at NHibernate.Cfg.Configuration.LogAndThrow(Exception exception) at NHibernate.Cfg.Configuration.AddInputStream(Stream xmlInputStream, String name) at NHibernate.Cfg.Configuration.AddResource(String path, Assembly assembly) at NHibernate.Cfg.Configuration.AddAssembly(Assembly assembly) at DomainModel.RepositoryBase..ctor() at WebController.Controllers._baseController..ctor() at WebController.Controllers.Admin.AdminProductController..ctor() at System.RuntimeType.CreateInstanceImpl(Boolean publicOnly, Boolean skipVisibilityChecks, Boolean fillCache) --- End of inner exception stack trace --- at System.RuntimeType.CreateInstanceImpl(Boolean publicOnly, Boolean skipVisibilityChecks, Boolean fillCache) at System.Activator.CreateInstance(Type type, Boolean nonPublic) at System.Web.Mvc.DefaultControllerFactory.GetControllerInstance(RequestContext requestContext, Type controllerType) --- End of inner exception stack trace --- at System.Web.Mvc.DefaultControllerFactory.GetControllerInstance(RequestContext requestContext, Type controllerType) at System.Web.Mvc.DefaultControllerFactory.CreateController(RequestContext requestContext, String controllerName) at System.Web.Mvc.MvcHandler.ProcessRequestInit(HttpContextBase httpContext, IController& controller, IControllerFactory& factory) at System.Web.Mvc.MvcHandler.BeginProcessRequest(HttpContextBase httpContext, AsyncCallback callback, Object state) at System.Web.HttpApplication.CallHandlerExecutionStep.System.Web.HttpApplication.IExecutionStep.Execute() at System.Web.HttpApplication.ExecuteStep(IExecutionStep step, Boolean& completedSynchronously) UPDATE CODE: In my Global.asax.cs, I'm doing this: protected void Application_BeginRequest(object sender, EventArgs e) { ManagedWebSessionContext.Bind(HttpContext.Current, SessionManager.SessionFactory.OpenSession()); } protected void Application_EndRequest(object sender, EventArgs e) { ISession session = ManagedWebSessionContext.Unbind(HttpContext.Current, SessionManager.SessionFactory); if (session != null) { try { if (session.Transaction != null && session.Transaction.IsActive) { session.Transaction.Rollback(); } else { session.Flush(); } } finally { session.Close(); } } } In the SessionManager class, I'm doing: public class SessionManager { private readonly ISessionFactory sessionFactory; public static ISessionFactory SessionFactory { get { return Instance.sessionFactory; } } private ISessionFactory GetSessionFactory() { return sessionFactory; } public static SessionManager Instance { get { return NestedSessionManager.sessionManager; } } public static ISession OpenSession() { return Instance.GetSessionFactory().OpenSession(); } public static ISession CurrentSession { get { return Instance.GetSessionFactory().GetCurrentSession(); } } private SessionManager() { Configuration config = new Configuration().Configure(); config.AddAssembly(Assembly.GetExecutingAssembly()); sessionFactory = config.BuildSessionFactory(); } class NestedSessionManager { internal static readonly SessionManager sessionManager = new SessionManager(); } } In the Repository, I'm doing this: public IEnumerable<User> GetAll() { ICriteria criteria = SessionManager.CurrentSession.CreateCriteria(typeof(User)); return criteria.List<User>(); } In the Controller, I'm doing this: public class UserController : _baseController { IUserRoleRepository _userRoleRepository; internal static readonly ILogger log = LogManager.GetLogger(typeof(UserController)); public UserController() { _userRoleRepository = new UserRoleRepository(); } public ActionResult UserList() { var myList = _usersRepository.GetAll(); return View(myList); } }

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• UITableView: Juxtaposing row, header, and footer insertions/deletions

  - by jdandrea

  Consider a very simple UITableView with one of two states. First state: One (overall) table footer One section containing two rows, a section header, and a section footer Second state: No table footer One section containing four rows and no section header/footer In both cases, each row is essentially one of four possible UITableViewCell objects, each containing its own UITextField. We don't even bother with reuse or caching, since we're only dealing with four known cells in this case. They've been created in an accompanying XIB, so we already have them all wired up and ready to go. Now consider we want to toggle between the two states. Sounds easy enough. Let's suppose our view controller's right bar button item provides the toggling support. We'll also track the current state with an ivar and enumeration. To be explicit for a sec, here's how one might go from state 1 to 2. (Presume we handle the bar button item's title as well.) In short, we want to clear out our table's footer view, then insert the third and fourth rows. We batch this inside an update block like so: // Brute forced references to the third and fourth rows in section 0 NSUInteger row02[] = {0, 2}; NSUInteger row03[] = {0, 3}; [self.tableView beginUpdates]; state = tableStateTwo; // 'internal' iVar, not a property self.tableView.tableFooterView = nil; [self.tableView insertRowsAtIndexPaths:[NSArray arrayWithObjects: [NSIndexPath indexPathWithIndexes:row02 length:2], [NSIndexPath indexPathWithIndexes:row03 length:2], nil] withRowAnimation:UITableViewRowAnimationFade]; [self.tableView endUpdates]; For the reverse, we want to reassign the table footer view (which, like the cells, is in the XIB ready and waiting), and remove the last two rows: // Use row02 and row03 from earlier snippet [self.tableView beginUpdates]; state = tableStateOne; self.tableView.tableFooterView = theTableFooterView; [self.tableView deleteRowsAtIndexPaths:[NSArray arrayWithObjects: [NSIndexPath indexPathWithIndexes:row02 length:2], [NSIndexPath indexPathWithIndexes:row03 length:2], nil] withRowAnimation:UITableViewRowAnimationFade]; [self.tableView endUpdates]; Now, when the table asks for rows, it's very cut and dry. The first two cells are the same in both cases. Only the last two appear/disappear depending on the state. The state ivar is consulted when the Table View asks for things like number of rows in a section, height for header/footer in a section, or view for header/footer in a section. This last bit is also where I'm running into trouble. Using the above logic, section 0's header/footer does not disappear. Specifically, the footer stays below the inserted rows, but the header now overlays the topmost row. If we switch back to state one, the section footer is removed, but the section header remains. How about using [self.tableView reloadData] then? Sure, why not. We take care not to use it inside the update block, per Apple's advisement, and simply add it after endUpdates. This time, good news! The section 0 header/footer disappears. :) However ... Toggling back to state one results in a most exquisite mess! The section 0 header returns, only to overlay the first row once again (instead of appear above it). The section 0 footer is placed below the last row just fine, but the overall table footer - now reinstated - overlays the section footer. Waaaaaah … now what? Just to be sure, let's toggle back to state two again. Yep, that looks fine. Coming back to state one? Yecccch. I also tried sprinkling in a few other stunts like using reloadSections:withRowAnimation:, but that only serves to make things worse. NSRange range = {0, 1}; NSIndexSet *indexSet = [NSIndexSet indexSetWithIndexesInRange:range]; ... [self.tableView reloadSections:indexSet withRowAnimation:UITableViewRowAnimationFade]; Case in point: If we invoke reloadSections... just before the end of the update block, changing to state two hides the first two rows from view, even though the space they would otherwise occupy remains. Switching back to state one returns section 0's header/footer to normal, but those first two rows remain invisible. Case two: Moving reloadSections... to just after the update block but before reloadData results in all rows becoming invisible! (I refer to the row as being invisible because, during tracing, tableView:cellForRowAtIndexPath: is returning bona-fide cell objects for those rows.) Case three: Moving reloadSections... after tableView:cellForRowAtIndexPath: brings us a bit closer, but the section 0 header/footer never returns when switching back to state one. Hmm. Perhaps it's a faux pas using both reloadSections... and reloadData, based on what I'm seeing at trace-time, which brings us to: Case four: Replacing reloadData with reloadSections... outright. All cells in state two disappear. All cells in state one remain missing as well (though the space is kept). So much for that theory. :) Tracing through the code, the cell and view objects, as well as the section heights, are all where they should be at the opportune times. They just aren't rendering sanely. So, how to crack this case? Clues welcome/appreciated!

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• Thread scheduling Round Robin / scheduling dispatch

  - by MRP

  #include <pthread.h> #include <stdio.h> #include <stdlib.h> #include <semaphore.h> #define NUM_THREADS 4 #define COUNT_LIMIT 13 int done = 0; int count = 0; int quantum = 2; int thread_ids[4] = {0,1,2,3}; int thread_runtime[4] = {0,5,4,7}; pthread_mutex_t count_mutex; pthread_cond_t count_threshold_cv; void * inc_count(void * arg); static sem_t count_sem; int quit = 0; ///////// Inc_Count//////////////// void *inc_count(void *t) { long my_id = (long)t; int i; sem_wait(&count_sem); /////////////CRIT SECTION////////////////////////////////// printf("run_thread = %d\n",my_id); printf("%d \n",thread_runtime[my_id]); for( i=0; i < thread_runtime[my_id];i++) { printf("runtime= %d\n",thread_runtime[my_id]); pthread_mutex_lock(&count_mutex); count++; if (count == COUNT_LIMIT) { pthread_cond_signal(&count_threshold_cv); printf("inc_count(): thread %ld, count = %d Threshold reached.\n", my_id, count); } printf("inc_count(): thread %ld, count = %d, unlocking mutex\n",my_id, count); pthread_mutex_unlock(&count_mutex); sleep(1) ; }//End For sem_post(&count_sem); // Next Thread Enters Crit Section pthread_exit(NULL); } /////////// Count_Watch //////////////// void *watch_count(void *t) { long my_id = (long)t; printf("Starting watch_count(): thread %ld\n", my_id); pthread_mutex_lock(&count_mutex); if (count<COUNT_LIMIT) { pthread_cond_wait(&count_threshold_cv, &count_mutex); printf("watch_count(): thread %ld Condition signal received.\n", my_id); printf("watch_count(): thread %ld count now = %d.\n", my_id, count); } pthread_mutex_unlock(&count_mutex); pthread_exit(NULL); } ////////////////// Main //////////////// int main (int argc, char *argv[]) { int i; long t1=0, t2=1, t3=2, t4=3; pthread_t threads[4]; pthread_attr_t attr; sem_init(&count_sem, 0, 1); /* Initialize mutex and condition variable objects */ pthread_mutex_init(&count_mutex, NULL); pthread_cond_init (&count_threshold_cv, NULL); /* For portability, explicitly create threads in a joinable state */ pthread_attr_init(&attr); pthread_attr_setdetachstate(&attr, PTHREAD_CREATE_JOINABLE); pthread_create(&threads[0], &attr, watch_count, (void *)t1); pthread_create(&threads[1], &attr, inc_count, (void *)t2); pthread_create(&threads[2], &attr, inc_count, (void *)t3); pthread_create(&threads[3], &attr, inc_count, (void *)t4); /* Wait for all threads to complete */ for (i=0; i<NUM_THREADS; i++) { pthread_join(threads[i], NULL); } printf ("Main(): Waited on %d threads. Done.\n", NUM_THREADS); /* Clean up and exit */ pthread_attr_destroy(&attr); pthread_mutex_destroy(&count_mutex); pthread_cond_destroy(&count_threshold_cv); pthread_exit(NULL); } I am trying to learn thread scheduling, there is a lot of technical coding that I don't know. I do know in theory how it should work, but having trouble getting started in code... I know, at least I think, this program is not real time and its not meant to be. Some how I need to create a scheduler dispatch to control the threads in the order they should run... RR FCFS SJF ect. Right now I don't have a dispatcher. What I do have is semaphores/ mutex to control the threads. This code does run FCFS... and I have been trying to use semaphores to create a RR.. but having a lot of trouble. I believe it would be easier to create a dispatcher but I dont know how. I need help, I am not looking for answers just direction.. some sample code will help to understand a bit more. Thank you.

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• Firefox throwing a exception with HTML Canvas putImageData

  - by mr.doob

  So I was working on this little javascript experiment and I needed a widget to track the FPS of it. I ported a widget I've been using with Actionscript 3 to Javascript and it seems to be working fine with Chrome/Safari but on Firefox is throwing an exception. This is the experiment: Depth of Field This is the error: [Exception... "An invalid or illegal string was specified" code: "12" nsresult: "0x8053000c (NS_ERROR_DOM_SYNTAX_ERR)" location: "http://mrdoob.com/projects/chromeexperiments/depth_of_field__debug/js/net/hires/debug/Stats.js Line: 105"] The line that is complaning about is this one: graph.putImageData(graphData, 1, 0, 0, 0, 69, 50); Which is a crappy code to "scroll" the bitmap pixels. The idea is that I only draw a few pixels on the left of the bitmap and then on the next frame I copy the whole bitmap and paste it on pixel to the right. This error usually is thrown because you're pasting a bitmap bigger than the source and it's going off the limits, but in theory that shouldn't be the case as I'm defining 69 as the width of the rectangle to paste (being the bitmap 70px wide). And this is full code: var Stats = { baseFps: null, timer: null, timerStart: null, timerLast: null, fps: null, ms: null, container: null, fpsText: null, msText: null, memText: null, memMaxText: null, graph: null, graphData: null, init: function(userfps) { baseFps = userfps; timer = 0; timerStart = new Date() - 0; timerLast = 0; fps = 0; ms = 0; container = document.createElement("div"); container.style.fontFamily = 'Arial'; container.style.fontSize = '10px'; container.style.backgroundColor = '#000033'; container.style.width = '70px'; container.style.paddingTop = '2px'; fpsText = document.createElement("div"); fpsText.style.color = '#ffff00'; fpsText.style.marginLeft = '3px'; fpsText.style.marginBottom = '-3px'; fpsText.innerHTML = "FPS:"; container.appendChild(fpsText); msText = document.createElement("div"); msText.style.color = '#00ff00'; msText.style.marginLeft = '3px'; msText.style.marginBottom = '-3px'; msText.innerHTML = "MS:"; container.appendChild(msText); memText = document.createElement("div"); memText.style.color = '#00ffff'; memText.style.marginLeft = '3px'; memText.style.marginBottom = '-3px'; memText.innerHTML = "MEM:"; container.appendChild(memText); memMaxText = document.createElement("div"); memMaxText.style.color = '#ff0070'; memMaxText.style.marginLeft = '3px'; memMaxText.style.marginBottom = '3px'; memMaxText.innerHTML = "MAX:"; container.appendChild(memMaxText); var canvas = document.createElement("canvas"); canvas.width = 70; canvas.height = 50; container.appendChild(canvas); graph = canvas.getContext("2d"); graph.fillStyle = '#000033'; graph.fillRect(0, 0, canvas.width, canvas.height ); graphData = graph.getImageData(0, 0, canvas.width, canvas.height); setInterval(this.update, 1000/baseFps); return container; }, update: function() { timer = new Date() - timerStart; if ((timer - 1000) > timerLast) { fpsText.innerHTML = "FPS: " + fps + " / " + baseFps; timerLast = timer; graph.putImageData(graphData, 1, 0, 0, 0, 69, 50); graph.fillRect(0,0,1,50); graphData = graph.getImageData(0, 0, 70, 50); var index = ( Math.floor(Math.min(50, (fps / baseFps) * 50)) * 280 /* 70 * 4 */ ); graphData.data[index] = graphData.data[index + 1] = 256; index = ( Math.floor(Math.min(50, 50 - (timer - ms) * .5)) * 280 /* 70 * 4 */ ); graphData.data[index + 1] = 256; graph.putImageData (graphData, 0, 0); fps = 0; } ++fps; msText.innerHTML = "MS: " + (timer - ms); ms = timer; } } Any ideas? Thanks in advance.

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• What strategy do you use for package naming in Java projects and why?

  - by Tim Visher

  I thought about this awhile ago and it recently resurfaced as my shop is doing its first real Java web app. As an intro, I see two main package naming strategies. (To be clear, I'm not referring to the whole 'domain.company.project' part of this, I'm talking about the package convention beneath that.) Anyway, the package naming conventions that I see are as follows: Functional: Naming your packages according to their function architecturally rather than their identity according to the business domain. Another term for this might be naming according to 'layer'. So, you'd have a *.ui package and a *.domain package and a *.orm package. Your packages are horizontal slices rather than vertical. This is much more common than logical naming. In fact, I don't believe I've ever seen or heard of a project that does this. This of course makes me leery (sort of like thinking that you've come up with a solution to an NP problem) as I'm not terribly smart and I assume everyone must have great reasons for doing it the way they do. On the other hand, I'm not opposed to people just missing the elephant in the room and I've never heard a an actual argument for doing package naming this way. It just seems to be the de facto standard. Logical: Naming your packages according to their business domain identity and putting every class that has to do with that vertical slice of functionality into that package. I have never seen or heard of this, as I mentioned before, but it makes a ton of sense to me. I tend to approach systems vertically rather than horizontally. I want to go in and develop the Order Processing system, not the data access layer. Obviously, there's a good chance that I'll touch the data access layer in the development of that system, but the point is that I don't think of it that way. What this means, of course, is that when I receive a change order or want to implement some new feature, it'd be nice to not have to go fishing around in a bunch of packages in order to find all the related classes. Instead, I just look in the X package because what I'm doing has to do with X. From a development standpoint, I see it as a major win to have your packages document your business domain rather than your architecture. I feel like the domain is almost always the part of the system that's harder to grok where as the system's architecture, especially at this point, is almost becoming mundane in its implementation. The fact that I can come to a system with this type of naming convention and instantly from the naming of the packages know that it deals with orders, customers, enterprises, products, etc. seems pretty darn handy. It seems like this would allow you to take much better advantage of Java's access modifiers. This allows you to much more cleanly define interfaces into subsystems rather than into layers of the system. So if you have an orders subsystem that you want to be transparently persistent, you could in theory just never let anything else know that it's persistent by not having to create public interfaces to its persistence classes in the dao layer and instead packaging the dao class in with only the classes it deals with. Obviously, if you wanted to expose this functionality, you could provide an interface for it or make it public. It just seems like you lose a lot of this by having a vertical slice of your system's features split across multiple packages. I suppose one disadvantage that I can see is that it does make ripping out layers a little bit more difficult. Instead of just deleting or renaming a package and then dropping a new one in place with an alternate technology, you have to go in and change all of the classes in all of the packages. However, I don't see this is a big deal. It may be from a lack of experience, but I have to imagine that the amount of times you swap out technologies pales in comparison to the amount of times you go in and edit vertical feature slices within your system. So I guess the question then would go out to you, how do you name your packages and why? Please understand that I don't necessarily think that I've stumbled onto the golden goose or something here. I'm pretty new to all this with mostly academic experience. However, I can't spot the holes in my reasoning so I'm hoping you all can so that I can move on. Thanks in advance!

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• Mixing inheritance mapping strategies in NHibernate

  - by MylesRip

  I have a rather large inheritance hierarchy in which some of the subclasses add very little and others add quite a bit. I don't want to map the entire hierarchy using either "table per class hierarchy" or "table per subclass" due to the size and complexity of the hierarchy. Ideally I'd like to mix mapping strategies such that portions of the hierarchy where the subclasses add very little are combined into a common table a la "table per class hierarchy" and subclasses that add a lot are broken out into a separate table. Using this approach, I would expect to have 2 or 3 tables with very little wasted space instead of either 1 table with lots of fields that don't apply to most of the objects, or 20+ tables, several of which would have only a couple of columns. In the NHibernate Reference Documentation version 2.1.0, I found section 8.1.4 "Mixing table per class hierarchy with table per subclass". This approach switches strategies partway down the hierarchy by using: ... <subclass ...> <join ...> <property ...> ... </join> </subclass> ... This is great in theory. In practice, though, I found that the schema was too restrictive in what was allowed inside the "join" element for me to be able to accomplish what I needed. Here is the related part of the schema definition: <xs:element name="join"> <xs:complexType> <xs:sequence> <xs:element ref="subselect" minOccurs="0" /> <xs:element ref="comment" minOccurs="0" /> <xs:element ref="key" /> <xs:choice minOccurs="0" maxOccurs="unbounded"> <xs:element ref="property" /> <xs:element ref="many-to-one" /> <xs:element ref="component" /> <xs:element ref="dynamic-component" /> <xs:element ref="any" /> <xs:element ref="map" /> <xs:element ref="set" /> <xs:element ref="list" /> <xs:element ref="bag" /> <xs:element ref="idbag" /> <xs:element ref="array" /> <xs:element ref="primitive-array" /> </xs:choice> <xs:element ref="sql-insert" minOccurs="0" /> <xs:element ref="sql-update" minOccurs="0" /> <xs:element ref="sql-delete" minOccurs="0" /> </xs:sequence> <xs:attribute name="table" use="required" type="xs:string" /> <xs:attribute name="schema" type="xs:string" /> <xs:attribute name="catalog" type="xs:string" /> <xs:attribute name="subselect" type="xs:string" /> <xs:attribute name="fetch" default="join"> <xs:simpleType> <xs:restriction base="xs:string"> <xs:enumeration value="join" /> <xs:enumeration value="select" /> </xs:restriction> </xs:simpleType> </xs:attribute> <xs:attribute name="inverse" default="false" type="xs:boolean"> </xs:attribute> <xs:attribute name="optional" default="false" type="xs:boolean"> </xs:attribute> </xs:complexType> </xs:element> As you can see, this allows the use of "property" child elements or "component" child elements, but not both. It also doesn't allow for "subclass" child elements to continue the hierarchy below the point at which the strategy was changed. Is there a way to accomplish this?

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• java runtime 6 with socks v5 proxy - Possible?

  - by rwired

  I have written an application that (amongst other things) runs a local service in windows that acts as a SOCKS v5 proxy for Firefox. I'm in the debugging phase right now and have found certain websites that don't work correctly. For example the Java Applet for Picture Uploading on Facebook.com fails because is is unable to lookup domains. My app overrides a hidden FF config setting network.proxy.socks__remote__dns setting it to true. The whole purpose of the app is to allow access to websites when behind a firewall (e.g. if the user is in China), so this setting is essential to ensure domains are resolved remotely also (and not just HTTP requests). In the JRE6 settings (documented here) there isn't an equivalent setting, and since remote DNS resolution is a feature of SOCKS v5 and not v4 as the documentation seems to imply I'm worried that it's just not possible. How can I programmatically make sure the JRE uses a SOCKS v5 proxy for all requests (including DNS)? UPDATE: Steps to reproduce this problem: Make sure you are behind a firewall that blocks (or redirects) internet access including DNS Install PuTTY and add a dynamic SSH tunnel on some port number of your choice (e.g. 9870). Then login to a remote server that has full access to the internet Launch Firefox and you will not be able to browse the web In FF network settings set the SOCKS v5 proxy to localhost:9870 In FF go to about:config, change network.proxy.socks__remote__dns to true You will now be able to browse the web. Go to facebook.com, login, go to your profile and attempt to use the picture uploader java applet to add some pictures It will fail with a series of class not found errors looking similar to: load: class com.facebook.facebookphotouploader5.FacebookPhotoUploader5.class not found. I believe this is failing because the JRE is unable to resolve the domain that the class resides on. I'm basing this belief on the fact that the documentation (http://java.sun.com/javase/6/docs/technotes/guides/deployment/deployment-guide/properties.html) talks only about SOCKS v4 (which as far as I know does not support remote DNS). My deployment.properties file is located in %APPDATA%\Sun\Java\Deployment. I can confirm that modifications I make in the Java Control Panel get written into that file. If instead of "Use browser setting" the network settings for Java I override and attempt to use the SOCKS proxy settings manually, I still have the issue. There does not seem to be an easy way to force the JRE to do DNS remotely through the Proxy. UPDATE 2: Without the SOCKS proxy, from my local client www.facebook.com resolves to 203.161.230.171 upload.facebook.com resolves to 64.33.88.161 Neither host is reachable (because of the firewall) If I login to the remote server, I get: www.facebook.com 69.63.187.17 upload.facebook.com 69.63.178.32 Both these IPs change after a few minutes, as it seems Facebook uses round-robin DNS and other load-balancing. With the Proxy settings set in Firefox, I can navigate to www.facebook.com without any difficulty (since DNS is being resolved remotely on the Proxy). Whey I go to the page with the Java applet it fails with the stacktrace messages I've already reported. However if I edit Windows\System32\drivers\etc\hosts, adding the correct IP for upload.facebook.com I can get the applet to load and work correctly (restart of FF is sometimes necessary). This evidence seems to support my theory that the Java Runtime is not resolving DNS on the Proxy, but instead just routing traffic though it. My application is for mass-deployment, and needs to work with java applets on other sites (not just facebook). I really need a work-around for this problem. UPDATE 3 Stacktrace dump a requested by ZZ Coder: load: class com.facebook.facebookphotouploader5.FacebookPhotoUploader5.class not found. java.lang.ClassNotFoundException: com.facebook.facebookphotouploader5.FacebookPhotoUploader5.class at sun.plugin2.applet.Applet2ClassLoader.findClass(Unknown Source) at java.lang.ClassLoader.loadClass(Unknown Source) at java.lang.ClassLoader.loadClass(Unknown Source) at sun.plugin2.applet.Plugin2ClassLoader.loadCode(Unknown Source) at sun.plugin2.applet.Plugin2Manager.createApplet(Unknown Source) at sun.plugin2.applet.Plugin2Manager$AppletExecutionRunnable.run(Unknown Source) at java.lang.Thread.run(Unknown Source) Caused by: java.net.SocketException: Connection reset at java.net.SocketInputStream.read(Unknown Source) at java.io.BufferedInputStream.fill(Unknown Source) at java.io.BufferedInputStream.read1(Unknown Source) at java.io.BufferedInputStream.read(Unknown Source) at sun.net.www.http.HttpClient.parseHTTPHeader(Unknown Source) at sun.net.www.http.HttpClient.parseHTTP(Unknown Source) at sun.net.www.http.HttpClient.parseHTTP(Unknown Source) at sun.net.www.protocol.http.HttpURLConnection.getInputStream(Unknown Source) at java.net.HttpURLConnection.getResponseCode(Unknown Source) at sun.plugin2.applet.Applet2ClassLoader.getBytes(Unknown Source) at sun.plugin2.applet.Applet2ClassLoader.access$000(Unknown Source) at sun.plugin2.applet.Applet2ClassLoader$1.run(Unknown Source) at java.security.AccessController.doPrivileged(Native Method) ... 7 more Exception: java.lang.ClassNotFoundException: com.facebook.facebookphotouploader5.FacebookPhotoUploader5.class Dumping class loader cache... Live entry: key=http://upload.facebook.com/controls/2008.10.10_v5.5.8/,FacebookPhotoUploader5.jar,FacebookPhotoUploader5.jar, refCount=1, threadGroup=sun.plugin2.applet.Applet2ThreadGroup[name=http://upload.facebook.com/controls/2008.10.10_v5.5.8/-threadGroup,maxpri=4] Done.

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• How to Stich to Image objects in Java

  - by Imran

  Hi, I have a scenario in which i`m getting a number of tiles (e.g.12) from my mapping server. Now for buffering and offline functions I need to join them all back again so that we have to deal with 1 single image object instead of 12. I ve tried to do it without JAI my code is below. package imagemerge; import java.awt.*; import java.awt.image.*; import java.awt.event.*; public class ImageSticher extends WindowAdapter { Image tile1; Image tile2; Image result; ColorModel colorModel; int width,height,widthr,heightr; //int t1,t2; int t12[]; public ImageSticher() { } public ImageSticher (Image img1,Image img2,int w,int h) { tile1=img1; tile2=img2; width=w; height=h; colorModel=ColorModel.getRGBdefault(); } public Image horizontalStich() throws Exception { widthr=width*2; heightr=height; t12=new int[widthr * heightr]; int t1[]=new int[width*height]; PixelGrabber p1 =new PixelGrabber(tile1, 0, 0, width, height, t1, 0, width); p1.grabPixels(); int t2[]=new int[width*height]; PixelGrabber p2 =new PixelGrabber(tile2, 0, 0, width, height, t1, 0, width); p2.grabPixels(); int y, x, rp, rpi; int red1, red2, redr; int green1, green2, greenr; int blue1, blue2, bluer; int alpha1, alpha2, alphar; for(y=0;y<heightr;y++) { for(x=0;x<widthr;x++) { //System.out.println(x); rpi=y*widthr+x; // index of resulting pixel; rp=0; //initializing resulting pixel System.out.println(rpi); if(x<(widthr/2)) // x is less than width , copy first tile { //System.out.println("tile1="+x); blue1 = t1[rpi] & 0x00ff; // ERROR occurs here green1=(t1[rpi] >> 8) & 0x00ff; red1=(t1[rpi] >> 16) & 0x00ff; alpha1 = (t1[rpi] >> 24) & 0x00ff; redr = (int)(red1 * 1.0); // copying red band pixel into redresult,,,,1.0 is the alpha valye redr = (redr < 0)?(0):((redr>255)?(255):(redr)); greenr = (int)(green1 * 1.0); // redr = (int)(red1 * 1.0); // greenr = (greenr < 0)?(0):((greenr>255)?(255):(greenr)); bluer = (int)(blue1 * 1.0); bluer = (bluer < 0)?(0):((bluer>255)?(255):(bluer)); alphar = 255; //resulting pixel computed rp = (((((alphar << 8) + (redr & 0x0ff)) << 8) + (greenr & 0x0ff)) << 8) + (bluer & 0x0ff); } else // index is ahead of half way...copy second tile { blue2 = t2[rpi] & 0x00ff; // blue band bit of first tile green2=(t2[rpi] >> 8) & 0x00ff; red2=(t2[rpi] >> 16) & 0x00ff; alpha2 = (t2[rpi] >> 24) & 0x00ff; redr = (int)(red2 * 1.0); // copying red band pixel into redresult,,,,1.0 is the alpha valye redr = (redr < 0)?(0):((redr>255)?(255):(redr)); greenr = (int)(green2 * 1.0); // redr = (int)(red2 * 1.0); // greenr = (greenr < 0)?(0):((greenr>255)?(255):(greenr)); bluer = (int)(blue2 * 1.0); bluer = (bluer < 0)?(0):((bluer>255)?(255):(bluer)); alphar = 255; //resulting pixel computed rp = (((((alphar << 8) + (redr & 0x0ff)) << 8) + (greenr & 0x0ff)) << 8) + (bluer & 0x0ff); } t12[rpi] = rp; // copying resulting pixel in the result int array which will be converted to image } } MemoryImageSource mis; if (t12!=null) { mis = new MemoryImageSource(widthr, heightr, colorModel, t12, 0, widthr); result = Toolkit.getDefaultToolkit().createImage(mis); return result; } return null; } } now to check the my theory Im trying to join or stich two tiles horizontaly but im getting the error : java.lang.ArrayIndexOutOfBoundsException: 90000 at imagemerge.ImageSticher.horizontalStich(ImageSticher.java:69) at imageStream.ImageStream.getImage(ImageStream.java:75) at imageStream.ImageStream.main(ImageStream.java:28) is there some kind of limitation because when stiching two images of 300 x 300 horizontally it means the resulting image will be 600 x 300 ... that would make 180000 index size but its giving error at 90000, what am I doing wrong here

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• Neural Network Always Produces Same/Similar Outputs for Any Input

  - by l33tnerd

  I have a problem where I am trying to create a neural network for Tic-Tac-Toe. However, for some reason, training the neural network causes it to produce nearly the same output for any given input. I did take a look at Artificial neural networks benchmark, but my network implementation is built for neurons with the same activation function for each neuron, i.e. no constant neurons. To make sure the problem wasn't just due to my choice of training set (1218 board states and moves generated by a genetic algorithm), I tried to train the network to reproduce XOR. The logistic activation function was used. Instead of using the derivative, I multiplied the error by output*(1-output) as some sources suggested that this was equivalent to using the derivative. I can put the Haskell source on HPaste, but it's a little embarrassing to look at. The network has 3 layers: the first layer has 2 inputs and 4 outputs, the second has 4 inputs and 1 output, and the third has 1 output. Increasing to 4 neurons in the second layer didn't help, and neither did increasing to 8 outputs in the first layer. I then calculated errors, network output, bias updates, and the weight updates by hand based on http://hebb.mit.edu/courses/9.641/2002/lectures/lecture04.pdf to make sure there wasn't an error in those parts of the code (there wasn't, but I will probably do it again just to make sure). Because I am using batch training, I did not multiply by x in equation (4) there. I am adding the weight change, though http://www.faqs.org/faqs/ai-faq/neural-nets/part2/section-2.html suggests to subtract it instead. The problem persisted, even in this simplified network. For example, these are the results after 500 epochs of batch training and of incremental training. Input |Target|Output (Batch) |Output(Incremental) [1.0,1.0]|[0.0] |[0.5003781562785173]|[0.5009731800870864] [1.0,0.0]|[1.0] |[0.5003740346965251]|[0.5006347214672715] [0.0,1.0]|[1.0] |[0.5003734471544522]|[0.500589332376345] [0.0,0.0]|[0.0] |[0.5003674110937019]|[0.500095157458231] Subtracting instead of adding produces the same problem, except everything is 0.99 something instead of 0.50 something. 5000 epochs produces the same result, except the batch-trained network returns exactly 0.5 for each case. (Heck, even 10,000 epochs didn't work for batch training.) Is there anything in general that could produce this behavior? Also, I looked at the intermediate errors for incremental training, and the although the inputs of the hidden/input layers varied, the error for the output neuron was always +/-0.12. For batch training, the errors were increasing, but extremely slowly and the errors were all extremely small (x10^-7). Different initial random weights and biases made no difference, either. Note that this is a school project, so hints/guides would be more helpful. Although reinventing the wheel and making my own network (in a language I don't know well!) was a horrible idea, I felt it would be more appropriate for a school project (so I know what's going on...in theory, at least. There doesn't seem to be a computer science teacher at my school). EDIT: Two layers, an input layer of 2 inputs to 8 outputs, and an output layer of 8 inputs to 1 output, produces much the same results: 0.5+/-0.2 (or so) for each training case. I'm also playing around with pyBrain, seeing if any network structure there will work. Edit 2: I am using a learning rate of 0.1. Sorry for forgetting about that. Edit 3: Pybrain's "trainUntilConvergence" doesn't get me a fully trained network, either, but 20000 epochs does, with 16 neurons in the hidden layer. 10000 epochs and 4 neurons, not so much, but close. So, in Haskell, with the input layer having 2 inputs & 2 outputs, hidden layer with 2 inputs and 8 outputs, and output layer with 8 inputs and 1 output...I get the same problem with 10000 epochs. And with 20000 epochs. Edit 4: I ran the network by hand again based on the MIT PDF above, and the values match, so the code should be correct unless I am misunderstanding those equations. Some of my source code is at http://hpaste.org/42453/neural_network__not_working; I'm working on cleaning my code somewhat and putting it in a Github (rather than a private Bitbucket) repository. All of the relevant source code is now at https://github.com/l33tnerd/hsann.

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• Problem measuring N times the execution time of a code block

  - by Nazgulled

  EDIT: I just found my problem after writing this long post explaining every little detail... If someone can give me a good answer on what I'm doing wrong and how can I get the execution time in seconds (using a float with 5 decimal places or so), I'll mark that as accepted. Hint: The problem was on how I interpreted the clock_getttime() man page. Hi, Let's say I have a function named myOperation that I need to measure the execution time of. To measure it, I'm using clock_gettime() as it was recommend here in one of the comments. My teacher recommends us to measure it N times so we can get an average, standard deviation and median for the final report. He also recommends us to execute myOperation M times instead of just one. If myOperation is a very fast operation, measuring it M times allow us to get a sense of the "real time" it takes; cause the clock being used might not have the required precision to measure such operation. So, execution myOperation only one time or M times really depends if the operation itself takes long enough for the clock precision we are using. I'm having trouble dealing with that M times execution. Increasing M decreases (a lot) the final average value. Which doesn't make sense to me. It's like this, on average you take 3 to 5 seconds to travel from point A to B. But then you go from A to B and back to A 5 times (which makes it 10 times, cause A to B is the same as B to A) and you measure that. Than you divide by 10, the average you get is supposed to be the same average you take traveling from point A to B, which is 3 to 5 seconds. This is what I want my code to do, but it's not working. If I keep increasing the number of times I go from A to B and back A, the average will be lower and lower each time, it makes no sense to me. Enough theory, here's my code: #include <stdio.h> #include <time.h> #define MEASUREMENTS 1 #define OPERATIONS 1 typedef struct timespec TimeClock; TimeClock diffTimeClock(TimeClock start, TimeClock end) { TimeClock aux; if((end.tv_nsec - start.tv_nsec) < 0) { aux.tv_sec = end.tv_sec - start.tv_sec - 1; aux.tv_nsec = 1E9 + end.tv_nsec - start.tv_nsec; } else { aux.tv_sec = end.tv_sec - start.tv_sec; aux.tv_nsec = end.tv_nsec - start.tv_nsec; } return aux; } int main(void) { TimeClock sTime, eTime, dTime; int i, j; for(i = 0; i < MEASUREMENTS; i++) { printf(" » MEASURE %02d\n", i+1); clock_gettime(CLOCK_REALTIME, &sTime); for(j = 0; j < OPERATIONS; j++) { myOperation(); } clock_gettime(CLOCK_REALTIME, &eTime); dTime = diffTimeClock(sTime, eTime); printf(" - NSEC (TOTAL): %ld\n", dTime.tv_nsec); printf(" - NSEC (OP): %ld\n\n", dTime.tv_nsec / OPERATIONS); } return 0; } Notes: The above diffTimeClock function is from this blog post. I replaced my real operation with myOperation() because it doesn't make any sense to post my real functions as I would have to post long blocks of code, you can easily code a myOperation() with whatever you like to compile the code if you wish. As you can see, OPERATIONS = 1 and the results are: » MEASURE 01 - NSEC (TOTAL): 27456580 - NSEC (OP): 27456580 For OPERATIONS = 100 the results are: » MEASURE 01 - NSEC (TOTAL): 218929736 - NSEC (OP): 2189297 For OPERATIONS = 1000 the results are: » MEASURE 01 - NSEC (TOTAL): 862834890 - NSEC (OP): 862834 For OPERATIONS = 10000 the results are: » MEASURE 01 - NSEC (TOTAL): 574133641 - NSEC (OP): 57413 Now, I'm not a math wiz, far from it actually, but this doesn't make any sense to me whatsoever. I've already talked about this with a friend that's on this project with me and he also can't understand the differences. I don't understand why the value is getting lower and lower when I increase OPERATIONS. The operation itself should take the same time (on average of course, not the exact same time), no matter how many times I execute it. You could tell me that that actually depends on the operation itself, the data being read and that some data could already be in the cache and bla bla, but I don't think that's the problem. In my case, myOperation is reading 5000 lines of text from an CSV file, separating the values by ; and inserting those values into a data structure. For each iteration, I'm destroying the data structure and initializing it again. Now that I think of it, I also that think that there's a problem measuring time with clock_gettime(), maybe I'm not using it right. I mean, look at the last example, where OPERATIONS = 10000. The total time it took was 574133641ns, which would be roughly 0,5s; that's impossible, it took a couple of minutes as I couldn't stand looking at the screen waiting and went to eat something.

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• How to implement Cocoa copyWithZone on derived object in MonoMac C#?

  - by Justin Aquadro

  I'm currently porting a small Winforms-based .NET application to use a native Mac front-end with MonoMac. The application has a TreeControl with icons and text, which does not exist out of the box in Cocoa. So far, I've ported almost all of the ImageAndTextCell code in Apple's DragNDrop example: https://developer.apple.com/library/mac/#samplecode/DragNDropOutlineView/Listings/ImageAndTextCell_m.html#//apple_ref/doc/uid/DTS40008831-ImageAndTextCell_m-DontLinkElementID_6, which is assigned to an NSOutlineView as a custom cell. It seems to be working almost perfectly, except that I have not figured out how to properly port the copyWithZone method. Unfortunately, this means the internal copies that NSOutlineView is making do not have the image field, and it leads to the images briefly vanishing during expand and collapse operations. The objective-c code in question is: - (id)copyWithZone:(NSZone *)zone { ImageAndTextCell *cell = (ImageAndTextCell *)[super copyWithZone:zone]; // The image ivar will be directly copied; we need to retain or copy it. cell->image = [image retain]; return cell; } The first line is what's tripping me up, as MonoMac does not expose a copyWithZone method, and I don't know how to otherwise call it. Update Based on current answers and additional research and testing, I've come up with a variety of models for copying an object. static List<ImageAndTextCell> _refPool = new List<ImageAndTextCell>(); // Method 1 static IntPtr selRetain = Selector.GetHandle ("retain"); [Export("copyWithZone:")] public virtual NSObject CopyWithZone(IntPtr zone) { ImageAndTextCell cell = new ImageAndTextCell() { Title = Title, Image = Image, }; Messaging.void_objc_msgSend (cell.Handle, selRetain); return cell; } // Method 2 [Export("copyWithZone:")] public virtual NSObject CopyWithZone(IntPtr zone) { ImageAndTextCell cell = new ImageAndTextCell() { Title = Title, Image = Image, }; _refPool.Add(cell); return cell; } [Export("dealloc")] public void Dealloc () { _refPool.Remove(this); this.Dispose(); } // Method 3 static IntPtr selRetain = Selector.GetHandle ("retain"); [Export("copyWithZone:")] public virtual NSObject CopyWithZone(IntPtr zone) { ImageAndTextCell cell = new ImageAndTextCell() { Title = Title, Image = Image, }; _refPool.Add(cell); Messaging.void_objc_msgSend (cell.Handle, selRetain); return cell; } // Method 4 static IntPtr selRetain = Selector.GetHandle ("retain"); static IntPtr selRetainCount = Selector.GetHandle("retainCount"); [Export("copyWithZone:")] public virtual NSObject CopyWithZone (IntPtr zone) { ImageAndTextCell cell = new ImageAndTextCell () { Title = Title, Image = Image, }; _refPool.Add (cell); Messaging.void_objc_msgSend (cell.Handle, selRetain); return cell; } public void PeriodicCleanup () { List<ImageAndTextCell> markedForDelete = new List<ImageAndTextCell> (); foreach (ImageAndTextCell cell in _refPool) { uint count = Messaging.UInt32_objc_msgSend (cell.Handle, selRetainCount); if (count == 1) markedForDelete.Add (cell); } foreach (ImageAndTextCell cell in markedForDelete) { _refPool.Remove (cell); cell.Dispose (); } } // Method 5 static IntPtr selCopyWithZone = Selector.GetHandle("copyWithZone:"); [Export("copyWithZone:")] public virtual NSObject CopyWithZone(IntPtr zone) { IntPtr copyHandle = Messaging.IntPtr_objc_msgSendSuper_IntPtr(SuperHandle, selCopyWithZone, zone); ImageAndTextCell cell = new ImageAndTextCell(copyHandle) { Image = Image, }; _refPool.Add(cell); return cell; } Method 1: Increases the retain count of the unmanaged object. The unmanaged object will persist persist forever (I think? dealloc never called), and the managed object will be harvested early. Seems to be lose-lose all-around, but runs in practice. Method 2: Saves a reference of the managed object. The unmanaged object is left alone, and dealloc appears to be invoked at a reasonable time by the caller. At this point the managed object is released and disposed. This seems reasonable, but on the downside the base type's dealloc won't be run (I think?) Method 3: Increases the retain count and saves a reference. Unmanaged and managed objects leak forever. Method 4: Extends Method 3 by adding a cleanup function that is run periodically (e.g. during Init of each new ImageAndTextCell object). The cleanup function checks the retain counts of the stored objects. A retain count of 1 means the caller has released it, so we should as well. Should eliminate leaking in theory. Method 5: Attempt to invoke the copyWithZone method on the base type, and then construct a new ImageAndTextView object with the resulting handle. Seems to do the right thing (the base data is cloned). Internally, NSObject bumps the retain count on objects constructed like this, so we also use the PeriodicCleanup function to release these objects when they're no longer used. Based on the above, I believe Method 5 is the best approach since it should be the only one that results in a truly correct copy of the base type data, but I don't know if the approach is inherently dangerous (I am also making some assumptions about the underlying implementation of NSObject). So far nothing bad has happened "yet", but if anyone is able to vet my analysis then I would be more confident going forward.

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• Trouble with copying dictionaries and using deepcopy on an SQLAlchemy ORM object

  - by Az

  Hi there, I'm doing a Simulated Annealing algorithm to optimise a given allocation of students and projects. This is language-agnostic pseudocode from Wikipedia: s ? s0; e ? E(s) // Initial state, energy. sbest ? s; ebest ? e // Initial "best" solution k ? 0 // Energy evaluation count. while k < kmax and e > emax // While time left & not good enough: snew ? neighbour(s) // Pick some neighbour. enew ? E(snew) // Compute its energy. if enew < ebest then // Is this a new best? sbest ? snew; ebest ? enew // Save 'new neighbour' to 'best found'. if P(e, enew, temp(k/kmax)) > random() then // Should we move to it? s ? snew; e ? enew // Yes, change state. k ? k + 1 // One more evaluation done return sbest // Return the best solution found. The following is an adaptation of the technique. My supervisor said the idea is fine in theory. First I pick up some allocation (i.e. an entire dictionary of students and their allocated projects, including the ranks for the projects) from entire set of randomised allocations, copy it and pass it to my function. Let's call this allocation aOld (it is a dictionary). aOld has a weight related to it called wOld. The weighting is described below. The function does the following: Let this allocation, aOld be the best_node From all the students, pick a random number of students and stick in a list Strip (DEALLOCATE) them of their projects ++ reflect the changes for projects (allocated parameter is now False) and lecturers (free up slots if one or more of their projects are no longer allocated) Randomise that list Try assigning (REALLOCATE) everyone in that list projects again Calculate the weight (add up ranks, rank 1 = 1, rank 2 = 2... and no project rank = 101) For this new allocation aNew, if the weight wNew is smaller than the allocation weight wOld I picked up at the beginning, then this is the best_node (as defined by the Simulated Annealing algorithm above). Apply the algorithm to aNew and continue. If wOld < wNew, then apply the algorithm to aOld again and continue. The allocations/data-points are expressed as "nodes" such that a node = (weight, allocation_dict, projects_dict, lecturers_dict) Right now, I can only perform this algorithm once, but I'll need to try for a number N (denoted by kmax in the Wikipedia snippet) and make sure I always have with me, the previous node and the best_node. So that I don't modify my original dictionaries (which I might want to reset to), I've done a shallow copy of the dictionaries. From what I've read in the docs, it seems that it only copies the references and since my dictionaries contain objects, changing the copied dictionary ends up changing the objects anyway. So I tried to use copy.deepcopy().These dictionaries refer to objects that have been mapped with SQLA. Questions: I've been given some solutions to the problems faced but due to my über green-ness with using Python, they all sound rather cryptic to me. Deepcopy isn't playing nicely with SQLA. I've been told thatdeepcopy on ORM objects probably has issues that prevent it from working as you'd expect. Apparently I'd be better off "building copy constructors, i.e. def copy(self): return FooBar(....)." Can someone please explain what that means? I checked and found out that deepcopy has issues because SQLAlchemy places extra information on your objects, i.e. an _sa_instance_state attribute, that I wouldn't want in the copy but is necessary for the object to have. I've been told: "There are ways to manually blow away the old _sa_instance_state and put a new one on the object, but the most straightforward is to make a new object with __init__() and set up the attributes that are significant, instead of doing a full deep copy." What exactly does that mean? Do I create a new, unmapped class similar to the old, mapped one? An alternate solution is that I'd have to "implement __deepcopy__() on your objects and ensure that a new _sa_instance_state is set up, there are functions in sqlalchemy.orm.attributes which can help with that." Once again this is beyond me so could someone kindly explain what it means? A more general question: given the above information are there any suggestions on how I can maintain the information/state for the best_node (which must always persist through my while loop) and the previous_node, if my actual objects (referenced by the dictionaries, therefore the nodes) are changing due to the deallocation/reallocation taking place? That is, without using copy?

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• SetWindowHookEx and execution blocking

  - by Kalaz

  Hello, I just wonder... I mainly use .NET but now I started to investigate WINAPI calls. For example I am using this piece of code to hook to the API functions. It starts freezing, when I try to debug the application... using System; using System.Diagnostics; using System.Runtime.InteropServices; using System.Threading; using System.Windows.Forms; public class Keyboard { private const int WH_KEYBOARD_LL = 13; private const int WM_KEYDOWN = 0x0100; private static LowLevelKeyboardProc _proc = HookCallback; private static IntPtr _hookID = IntPtr.Zero; public static event Action<Keys,bool, bool> KeyDown; public static void Hook() { new Thread(new ThreadStart(()=> { _hookID = SetHook(_proc); Application.Run(); })).Start(); } public static void Unhook() { UnhookWindowsHookEx(_hookID); } private static IntPtr SetHook(LowLevelKeyboardProc proc) { using (Process curProcess = Process.GetCurrentProcess()) using (ProcessModule curModule = curProcess.MainModule) { return SetWindowsHookEx(WH_KEYBOARD_LL, proc, GetModuleHandle(curModule.ModuleName), 0); } } private delegate IntPtr LowLevelKeyboardProc( int nCode, IntPtr wParam, IntPtr lParam); private static IntPtr HookCallback( int nCode, IntPtr wParam, IntPtr lParam) { if (nCode >= 0 && wParam == (IntPtr)WM_KEYDOWN) { int vkCode = Marshal.ReadInt32(lParam); Keys k = (Keys) vkCode; if (KeyDown != null) { KeyDown.BeginInvoke(k, IsKeyPressed(VirtualKeyStates.VK_CONTROL), IsKeyPressed(VirtualKeyStates.VK_SHIFT),null,null); } } return CallNextHookEx(_hookID, nCode, wParam, lParam); } private static bool IsKeyPressed(VirtualKeyStates virtualKeyStates) { return (GetKeyState(virtualKeyStates) & (1 << 7))==128; } [DllImport("user32.dll", CharSet = CharSet.Auto, SetLastError = true)] private static extern IntPtr SetWindowsHookEx(int idHook, LowLevelKeyboardProc lpfn, IntPtr hMod, uint dwThreadId); [DllImport("user32.dll", CharSet = CharSet.Auto, SetLastError = true)] [return: MarshalAs(UnmanagedType.Bool)] private static extern bool UnhookWindowsHookEx(IntPtr hhk); [DllImport("user32.dll", CharSet = CharSet.Auto, SetLastError = true)] private static extern IntPtr CallNextHookEx(IntPtr hhk, int nCode, IntPtr wParam, IntPtr lParam); [DllImport("kernel32.dll", CharSet = CharSet.Auto, SetLastError = true)] private static extern IntPtr GetModuleHandle(string lpModuleName); [DllImport("user32.dll")] static extern short GetKeyState(VirtualKeyStates nVirtKey); } enum VirtualKeyStates : int { VK_LBUTTON = 0x01, VK_RBUTTON = 0x02, VK_CANCEL = 0x03, VK_MBUTTON = 0x04, // VK_XBUTTON1 = 0x05, VK_XBUTTON2 = 0x06, // VK_BACK = 0x08, VK_TAB = 0x09, // VK_CLEAR = 0x0C, VK_RETURN = 0x0D, // VK_SHIFT = 0x10, VK_CONTROL = 0x11, VK_MENU = 0x12, VK_PAUSE = 0x13, VK_CAPITAL = 0x14, // VK_KANA = 0x15, VK_HANGEUL = 0x15, /* old name - should be here for compatibility */ VK_HANGUL = 0x15, VK_JUNJA = 0x17, VK_FINAL = 0x18, VK_HANJA = 0x19, VK_KANJI = 0x19, // VK_ESCAPE = 0x1B, // VK_CONVERT = 0x1C, VK_NONCONVERT = 0x1D, VK_ACCEPT = 0x1E, VK_MODECHANGE = 0x1F, // VK_SPACE = 0x20, VK_PRIOR = 0x21, VK_NEXT = 0x22, VK_END = 0x23, VK_HOME = 0x24, VK_LEFT = 0x25, VK_UP = 0x26, VK_RIGHT = 0x27, VK_DOWN = 0x28, VK_SELECT = 0x29, VK_PRINT = 0x2A, VK_EXECUTE = 0x2B, VK_SNAPSHOT = 0x2C, VK_INSERT = 0x2D, VK_DELETE = 0x2E, VK_HELP = 0x2F, // VK_LWIN = 0x5B, VK_RWIN = 0x5C, VK_APPS = 0x5D, // VK_SLEEP = 0x5F, // VK_NUMPAD0 = 0x60, VK_NUMPAD1 = 0x61, VK_NUMPAD2 = 0x62, VK_NUMPAD3 = 0x63, VK_NUMPAD4 = 0x64, VK_NUMPAD5 = 0x65, VK_NUMPAD6 = 0x66, VK_NUMPAD7 = 0x67, VK_NUMPAD8 = 0x68, VK_NUMPAD9 = 0x69, VK_MULTIPLY = 0x6A, VK_ADD = 0x6B, VK_SEPARATOR = 0x6C, VK_SUBTRACT = 0x6D, VK_DECIMAL = 0x6E, VK_DIVIDE = 0x6F, VK_F1 = 0x70, VK_F2 = 0x71, VK_F3 = 0x72, VK_F4 = 0x73, VK_F5 = 0x74, VK_F6 = 0x75, VK_F7 = 0x76, VK_F8 = 0x77, VK_F9 = 0x78, VK_F10 = 0x79, VK_F11 = 0x7A, VK_F12 = 0x7B, VK_F13 = 0x7C, VK_F14 = 0x7D, VK_F15 = 0x7E, VK_F16 = 0x7F, VK_F17 = 0x80, VK_F18 = 0x81, VK_F19 = 0x82, VK_F20 = 0x83, VK_F21 = 0x84, VK_F22 = 0x85, VK_F23 = 0x86, VK_F24 = 0x87, // VK_NUMLOCK = 0x90, VK_SCROLL = 0x91, // VK_OEM_NEC_EQUAL = 0x92, // '=' key on numpad // VK_OEM_FJ_JISHO = 0x92, // 'Dictionary' key VK_OEM_FJ_MASSHOU = 0x93, // 'Unregister word' key VK_OEM_FJ_TOUROKU = 0x94, // 'Register word' key VK_OEM_FJ_LOYA = 0x95, // 'Left OYAYUBI' key VK_OEM_FJ_ROYA = 0x96, // 'Right OYAYUBI' key // VK_LSHIFT = 0xA0, VK_RSHIFT = 0xA1, VK_LCONTROL = 0xA2, VK_RCONTROL = 0xA3, VK_LMENU = 0xA4, VK_RMENU = 0xA5, // VK_BROWSER_BACK = 0xA6, VK_BROWSER_FORWARD = 0xA7, VK_BROWSER_REFRESH = 0xA8, VK_BROWSER_STOP = 0xA9, VK_BROWSER_SEARCH = 0xAA, VK_BROWSER_FAVORITES = 0xAB, VK_BROWSER_HOME = 0xAC, // VK_VOLUME_MUTE = 0xAD, VK_VOLUME_DOWN = 0xAE, VK_VOLUME_UP = 0xAF, VK_MEDIA_NEXT_TRACK = 0xB0, VK_MEDIA_PREV_TRACK = 0xB1, VK_MEDIA_STOP = 0xB2, VK_MEDIA_PLAY_PAUSE = 0xB3, VK_LAUNCH_MAIL = 0xB4, VK_LAUNCH_MEDIA_SELECT = 0xB5, VK_LAUNCH_APP1 = 0xB6, VK_LAUNCH_APP2 = 0xB7, // VK_OEM_1 = 0xBA, // ';:' for US VK_OEM_PLUS = 0xBB, // '+' any country VK_OEM_COMMA = 0xBC, // ',' any country VK_OEM_MINUS = 0xBD, // '-' any country VK_OEM_PERIOD = 0xBE, // '.' any country VK_OEM_2 = 0xBF, // '/?' for US VK_OEM_3 = 0xC0, // '`~' for US // VK_OEM_4 = 0xDB, // '[{' for US VK_OEM_5 = 0xDC, // '\|' for US VK_OEM_6 = 0xDD, // ']}' for US VK_OEM_7 = 0xDE, // ''"' for US VK_OEM_8 = 0xDF, // VK_OEM_AX = 0xE1, // 'AX' key on Japanese AX kbd VK_OEM_102 = 0xE2, // "<>" or "\|" on RT 102-key kbd. VK_ICO_HELP = 0xE3, // Help key on ICO VK_ICO_00 = 0xE4, // 00 key on ICO // VK_PROCESSKEY = 0xE5, // VK_ICO_CLEAR = 0xE6, // VK_PACKET = 0xE7, // VK_OEM_RESET = 0xE9, VK_OEM_JUMP = 0xEA, VK_OEM_PA1 = 0xEB, VK_OEM_PA2 = 0xEC, VK_OEM_PA3 = 0xED, VK_OEM_WSCTRL = 0xEE, VK_OEM_CUSEL = 0xEF, VK_OEM_ATTN = 0xF0, VK_OEM_FINISH = 0xF1, VK_OEM_COPY = 0xF2, VK_OEM_AUTO = 0xF3, VK_OEM_ENLW = 0xF4, VK_OEM_BACKTAB = 0xF5, // VK_ATTN = 0xF6, VK_CRSEL = 0xF7, VK_EXSEL = 0xF8, VK_EREOF = 0xF9, VK_PLAY = 0xFA, VK_ZOOM = 0xFB, VK_NONAME = 0xFC, VK_PA1 = 0xFD, VK_OEM_CLEAR = 0xFE } It works well even if you put messagebox into the event or something that blocks execution. But it gets bad if you try to put breakpoint into the event. Why? I mean event is not run in the same thread that the windows hook is. That means that It shouldn't block HookCallback. It does however... I would really like to know why is this happening. My theory is that Visual Studio when breaking execution temporarily stops all threads and that means that HookCallback is blocked... Is there any book or valuable resource that would explain concepts behind all of this threading?

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• The impossible inline Javascript delay/sleep

  - by trex005

  There is a JavaScript function, of which I have zero control of the code, which calls a function that I wrote. My function uses DOM to generate an iFrame, defines it's src and then appends it to another DOM element. However, before my function returns, and thus allows continued execution of the containing function, it is imperative that the iFrame be fully loaded. Here are the things that I have tried and why they do not work : 1. The SetTimeout option : 99.999% of the time, this is THE answer. As a matter of fact, in the past decade that I have been mentoring in JavaScript, I have always insisted that code could always be refactored to use this option, and never believed a scenario existed where that was not the case. Well, I finally found one! The problem is that because my function is being called inline, if the very next line is executed before my iFrame finishes loading, it totally neuters my script, and since the moment my script completes, the external script continues. A callback of sorts will not work 2. The "Do nothing" loop :This option you use while(//iFrame is not loaded){//do nothing}. In theory this would not return until the frame is loaded. The problem is that since this hogs all the resources, the iFrame never loads. This trick, although horribly unprofessional, dirty etc. will work when you just need an inline delay, but since I require an external thread to complete, it will not.In FF, after a few seconds, it pauses the script and an alert pops up stating that there is an unresponsive script. While that alert is up, the iFrame is able to load, and then my function is able to return, but having the browser frozen for 10 seconds, and then requiring the user to correctly dismiss an error is a no go. 3. The model dialogue : I was inspired by the fact that the FF popup allowed the iFrame to load while halting the execution of the function, and thinking about it, I realized that it is because the modal dialogue, is a way of halting execution yet allowing other threads to continue! Brilliant, so I decided to try other modal options. Things like alert() work beautifully! When it pops up, even if only up for 1/10th of a second, the iFrame is able to complete, and all works great. And just in case the 1/10 of a second is not sufficient, I can put the model dialogue in the while loop from solution 2, and it would ensure that the iFrame is loaded in time. Sweet right? Except for the fact that I now have to pop up a very unprofessional dialogue for the user to dismiss in order to run my script. I fought with myself about this cost/benefit of this action, but then I encountered a scenario where my code was called 10 times on a single page! Having to dismiss 10 alerts before acessing a page?! That reminds me of the late 90s script kiddie pages, and is NOT an option. 4. A gazillion other delay script out there:There are about 10 jQuery delay or sleep functions, some of them actually quite cleverly developed, but none worked. A few prototype options, and again, none I found could do it! A dozen or so other libraries and frameworks claimed they had what I needed, but alas they all conspired to give me false hope. I am convinced that since a built in model dialogue can halt execution, while allowing other threads to continue, there must be some code accessible way to do the same thing with out user input. The Code is literally thousands upon thousands of lines and is proprietary, so I wrote this little example of the problem for you to work with. It is important to note the ONLY code you are able to change is in the onlyThingYouCanChange function Test File : <html> <head> </head> </html> <body> <div id='iFrameHolder'></div> <script type='text/javascript'> function unChangeableFunction() { new_iFrame = onlyThingYouCanChange(document.getElementById('iFrameHolder')); new_iFrame_doc = (new_iFrame.contentWindow || new_iFrame.contentDocument); if(new_iFrame_doc.document)new_iFrame_doc=new_iFrame_doc.document; new_iFrame_body = new_iFrame_doc.body; if(new_iFrame_body.innerHTML != 'Loaded?') { //The world explodes!!! alert('you just blew up the world! Way to go!'); } else { alert('wow, you did it! Way to go!'); } } var iFrameLoaded = false; function onlyThingYouCanChange(objectToAppendIFrameTo) { iFrameLoaded = false; iframe=document.createElement('iframe'); iframe.onload = new Function('iFrameLoaded = true'); iframe.src = 'blank_frame.html'; //Must use an HTML doc on the server because there is a very specific DOM structure that must be maintained. objectToAppendIFrameTo.appendChild(iframe); var it = 0; while(!iFrameLoaded) //I put the limit on here so you don't { //If I was able to put some sort of delay here that paused the exicution of the script, but did not halt all other browser threads, and did not require user interaction we'd be golden! //alert('test'); //This would work if it did not require user interaction! } return iframe; } unChangeableFunction(); </script> </body> blank_frame.html : <html> <head> </head> <body style='margin:0px'>Loaded?</body> </html>

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• timer_getoverrun() doesn't behave as expected when using sleep()

  - by dlp

  Here is a program that uses a POSIX per-process timer alongside the sleep subroutine. The signal used by the timer has been set to SIGUSR1 rather than SIGALRM, since SIGALRM may be used internally by sleep, but it still doesn't seem to work. I have run the program using the command line timer-overruns -d 1 -n 10000000 (1 cs interval) so, in theory, we should expect 100 overruns between calls to sigwaitinfo. However, timer_getoverrun returns 0. I have also tried a version using a time-consuming for loop to introduce the delay. In this case, overruns are recorded. Does anyone know why this happens? I am running a 3.4 Linux kernel. Program source /* * timer-overruns.c */ #include <unistd.h> #include <stdlib.h> #include <stdio.h> #include <signal.h> #include <time.h> // Signal to be used for timer expirations #define TIMER_SIGNAL SIGUSR1 int main(int argc, char **argv) { int opt; int d = 0; int r = 0; // Repeat indefinitely struct itimerspec its; its.it_interval.tv_sec = 0; its.it_interval.tv_nsec = 0; // Parse arguments while ((opt = getopt(argc, argv, "d:r:s:n:")) != -1) { switch (opt) { case 'd': // Delay before calling sigwaitinfo() d = atoi(optarg); break; case 'r': // Number of times to call sigwaitinfo() r = atoi(optarg); break; case 's': // Timer interval (seconds) its.it_interval.tv_sec = its.it_value.tv_sec = atoi(optarg); break; case 'n': // Timer interval (nanoseconds) its.it_interval.tv_nsec = its.it_value.tv_nsec = atoi(optarg); break; default: /* '?' */ fprintf(stderr, "Usage: %s [-d signal_accept_delay] [-r repetitions] [-s interval_seconds] [-n interval_nanoseconds]\n", argv[0]); exit(EXIT_FAILURE); } } // Check sanity of command line arguments short e = 0; if (d < 0) { fprintf(stderr, "Delay (-d) cannot be negative!\n"); e++; } if (r < 0) { fprintf(stderr, "Number of repetitions (-r) cannot be negative!\n"); e++; } if (its.it_interval.tv_sec < 0) { fprintf(stderr, "Interval seconds value (-s) cannot be negative!\n"); e++; } if (its.it_interval.tv_nsec < 0) { fprintf(stderr, "Interval nanoseconds value (-n) cannot be negative!\n"); e++; } if (its.it_interval.tv_nsec > 999999999) { fprintf(stderr, "Interval nanoseconds value (-n) must be < 1 second.\n"); e++; } if (e > 0) exit(EXIT_FAILURE); // Set default values if not specified if (its.it_interval.tv_sec == 0 && its.it_interval.tv_nsec == 0) { its.it_interval.tv_sec = its.it_value.tv_sec = 1; its.it_value.tv_nsec = 0; } printf("Running with timer delay %d.%09d seconds\n", (int) its.it_interval.tv_sec, (int) its.it_interval.tv_nsec); // Will be waiting for signals synchronously, so block the one in use. sigset_t sigset; sigemptyset(&sigset); sigaddset(&sigset, TIMER_SIGNAL); sigprocmask(SIG_BLOCK, &sigset, NULL ); // Create and arm the timer struct sigevent sev; timer_t timer; sev.sigev_notify = SIGEV_SIGNAL; sev.sigev_signo = TIMER_SIGNAL; sev.sigev_value.sival_ptr = timer; timer_create(CLOCK_REALTIME, &sev, &timer); timer_settime(timer, TIMER_ABSTIME, &its, NULL ); // Signal handling loop int overruns; siginfo_t si; // Make the loop infinite if r = 0 if (r == 0) r = -1; while (r != 0) { // Sleeping should cause overruns if (d > 0) sleep(d); sigwaitinfo(&sigset, &si); // Check that the signal is from the timer if (si.si_code != SI_TIMER) continue; overruns = timer_getoverrun(timer); if (overruns > 0) { printf("Timer overrun occurred for %d expirations.\n", overruns); } // Decrement r if not repeating indefinitely if (r > 0) r--; } return EXIT_SUCCESS; }

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• bandwidth throttling C linux

  - by bob moch

  hi im currently creating a function to create a sleep time i can pause between packets for my port scanner im creating for personal/educational use for my home network. what im currently doing is opening /proc/net/dev and reading the 9th set of digits for the eth0 interface to find out the current packets being set and then reading it again and doing some math to figure out a delay to sleep between sending a packet to a port to identify it and fingerprint it. my problem is that no matter what throttle % i use it always seems to send the same rate of packets. i think its mainly my way of mathematically creating my sleep delay. edit:: dont mind the function declaration and the struct stuff all im doing is spawning this function in a thread and passing a pointer to a struct to the function, recreating the struct locally and then freeing the passed structs memory. void *bandwidthmonitor_cmd(void *param) { char cmdline[1024], *bytedata[19]; int i = 0, ii = 0; long long prevbytes = 0, currentbytes = 0, elapsedbytes = 0, byteusage = 0, maxthrottle = 0; command_struct bandwidth = *((command_struct *)param); free(param); //printf("speed: %d\n throttle: %d\n\n", UPLOAD_SPEED, bandwidth.throttle); maxthrottle = UPLOAD_SPEED * bandwidth.throttle / 100; //printf("max throttle:%lld\n", maxthrottle); FILE *f = fopen("/proc/net/dev", "r"); if(f != NULL) { while(1) { while(fgets(cmdline, sizeof(cmdline), f) != NULL) { cmdline[strlen(cmdline)] = '\0'; if(strncmp(cmdline, " eth0", 6) == 0) { bytedata[0] = strtok(cmdline, " "); while(bytedata[i] != NULL) { i++; bytedata[i] = strtok(NULL, " "); } bytedata[i + 1] = '\0'; currentbytes = atoi(bytedata[9]); } } i = 0; rewind(f); elapsedbytes = currentbytes - prevbytes; prevbytes = currentbytes; byteusage = 8 * (elapsedbytes / 1024); //printf("usage:%lld\n",byteusage); if(ii & 0x40) { SLEEP += (maxthrottle - byteusage) * -1.1;//-2.5; if(SLEEP < 0){ SLEEP = 0; } //printf("sleep:%d\n", SLEEP); } usleep(25000); ii++; } } return NULL; } SLEEP and UPLOAD_SPEED are global variables and UPLOAD_SPEED is in kb/s and generated via a speedtest function that gets the upload speed of my computer. this function is running inside a POSIX thread updating SLEEP which my threads doing the socket work grab to sleep by after every packet. as testing instead of only doing the ports i want to check i make it do all the ports over and over again so i can run dstat on a machine to check bandwidth and no matter what bandwidth.throttle is set to it always seems to generate the same amount of bandwidth to the dstat machine. the way i calculate how much i "should" throttle by is by finding the maximum throttle speed which is defined as maxthrottle = upload_speed * throttle / 100; for example if my upload speed was 1000kb/s and my throttle was 90 (90%) my max throttle would be 900kb/s from there it would find the current bytes sent from /proc/net/dev and then find my sleep time via incrementing or decrementing it via sleep += (maxthrottle - bytesysed) * -1.1; this should in theory increase or decrease the sleep time based on how many bytes used there are. the if(ii & 0x40) statement is just for some moderation control. it makes it so it only sets sleep to a new time every 30-40 iterations. final notes: the main problem is that the sleep timer does not seem to modify the speed of packets being set. or maybe its just my implementation because on a freshly restarted machine where /proc/net/dev has low numbers of bytes sent it seems to raise the sleep timer accordingly on my 60kb/s upload machine (ex if i set the throttle to 2 it will incline the sleep timer until network bandwidth out reaches the max bandwidth threshold, but when i try running it on a server which as been online forever it doesnt seem to work as nicely if at all. if anyone can suggest a new method of monitoring the network to adjust a sleep delay then let me know or if anyone sees a flaw in my code. thank you.

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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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• Why is Java EE 6 better than Spring ?

  - by arungupta

  Java EE 6 was released over 2 years ago and now there are 14 compliant application servers. In all my talks around the world, a question that is frequently asked is Why should I use Java EE 6 instead of Spring ? There are already several blogs covering that topic: Java EE wins over Spring by Bill Burke Why will I use Java EE instead of Spring in new Enterprise Java projects in 2012 ? by Kai Waehner (more discussion on TSS) Spring to Java EE migration (Part 1 and 2, 3 and 4 coming as well) by David Heffelfinger Spring to Java EE - A Migration Experience by Lincoln Baxter Migrating Spring to Java EE 6 by Bert Ertman and Paul Bakker at NLJUG Moving from Spring to Java EE 6 - The Age of Frameworks is Over at TSS Java EE vs Spring Shootout by Rohit Kelapure and Reza Rehman at JavaOne 2011 Java EE 6 and the Ewoks by Murat Yener Definite excuse to avoid Spring forever - Bert Ertman and Arun Gupta I will try to share my perspective in this blog. First of all, I'd like to start with a note: Thank you Spring framework for filling the interim gap and providing functionality that is now included in the mainstream Java EE 6 application servers. The Java EE platform has evolved over the years learning from frameworks like Spring and provides all the functionality to build an enterprise application. Thank you very much Spring framework! While Spring was revolutionary in its time and is still very popular and quite main stream in the same way Struts was circa 2003, it really is last generation's framework - some people are even calling it legacy. However my theory is "code is king". So my approach is to build/take a simple Hello World CRUD application in Java EE 6 and Spring and compare the deployable artifacts. I started looking at the official tutorial Developing a Spring Framework MVC Application Step-by-Step but it is using the older version 2.5. I wasn't able to find any updated version in the current 3.1 release. Next, I downloaded Spring Tool Suite and thought that would provide some template samples to get started. A least a quick search did not show any handy tutorials - either video or text-based. So I searched and found a link to their SVN repository at src.springframework.org/svn/spring-samples/. I tried the "mvc-basic" sample and the generated WAR file was 4.43 MB. While it was named a "basic" sample it seemed to come with 19 different libraries bundled but it was what I could find: ./WEB-INF/lib/aopalliance-1.0.jar./WEB-INF/lib/hibernate-validator-4.1.0.Final.jar./WEB-INF/lib/jcl-over-slf4j-1.6.1.jar./WEB-INF/lib/joda-time-1.6.2.jar./WEB-INF/lib/joda-time-jsptags-1.0.2.jar./WEB-INF/lib/jstl-1.2.jar./WEB-INF/lib/log4j-1.2.16.jar./WEB-INF/lib/slf4j-api-1.6.1.jar./WEB-INF/lib/slf4j-log4j12-1.6.1.jar./WEB-INF/lib/spring-aop-3.0.5.RELEASE.jar./WEB-INF/lib/spring-asm-3.0.5.RELEASE.jar./WEB-INF/lib/spring-beans-3.0.5.RELEASE.jar./WEB-INF/lib/spring-context-3.0.5.RELEASE.jar./WEB-INF/lib/spring-context-support-3.0.5.RELEASE.jar./WEB-INF/lib/spring-core-3.0.5.RELEASE.jar./WEB-INF/lib/spring-expression-3.0.5.RELEASE.jar./WEB-INF/lib/spring-web-3.0.5.RELEASE.jar./WEB-INF/lib/spring-webmvc-3.0.5.RELEASE.jar./WEB-INF/lib/validation-api-1.0.0.GA.jar And it is not even using any database! The app deployed fine on GlassFish 3.1.2 but the "@Controller Example" link did not work as it was missing the context root. With a bit of tweaking I could deploy the application and assume that the account got created because no error was displayed in the browser or server log. Next I generated the WAR for "mvc-ajax" and the 5.1 MB WAR had 20 JARs (1 removed, 2 added): ./WEB-INF/lib/aopalliance-1.0.jar./WEB-INF/lib/hibernate-validator-4.1.0.Final.jar./WEB-INF/lib/jackson-core-asl-1.6.4.jar./WEB-INF/lib/jackson-mapper-asl-1.6.4.jar./WEB-INF/lib/jcl-over-slf4j-1.6.1.jar./WEB-INF/lib/joda-time-1.6.2.jar./WEB-INF/lib/jstl-1.2.jar./WEB-INF/lib/log4j-1.2.16.jar./WEB-INF/lib/slf4j-api-1.6.1.jar./WEB-INF/lib/slf4j-log4j12-1.6.1.jar./WEB-INF/lib/spring-aop-3.0.5.RELEASE.jar./WEB-INF/lib/spring-asm-3.0.5.RELEASE.jar./WEB-INF/lib/spring-beans-3.0.5.RELEASE.jar./WEB-INF/lib/spring-context-3.0.5.RELEASE.jar./WEB-INF/lib/spring-context-support-3.0.5.RELEASE.jar./WEB-INF/lib/spring-core-3.0.5.RELEASE.jar./WEB-INF/lib/spring-expression-3.0.5.RELEASE.jar./WEB-INF/lib/spring-web-3.0.5.RELEASE.jar./WEB-INF/lib/spring-webmvc-3.0.5.RELEASE.jar./WEB-INF/lib/validation-api-1.0.0.GA.jar 2 more JARs for just doing Ajax. Anyway, deploying this application gave the following error: Caused by: java.lang.NoSuchMethodError: org.codehaus.jackson.map.SerializationConfig.<init>(Lorg/codehaus/jackson/map/ClassIntrospector;Lorg/codehaus/jackson/map/AnnotationIntrospector;Lorg/codehaus/jackson/map/introspect/VisibilityChecker;Lorg/codehaus/jackson/map/jsontype/SubtypeResolver;)V    at org.springframework.samples.mvc.ajax.json.ConversionServiceAwareObjectMapper.<init>(ConversionServiceAwareObjectMapper.java:20)    at org.springframework.samples.mvc.ajax.json.JacksonConversionServiceConfigurer.postProcessAfterInitialization(JacksonConversionServiceConfigurer.java:40)    at org.springframework.beans.factory.support.AbstractAutowireCapableBeanFactory.applyBeanPostProcessorsAfterInitialization(AbstractAutowireCapableBeanFactory.java:407) Seems like some incorrect repos in the "pom.xml". Next one is "mvc-showcase" and the 6.49 MB WAR now has 28 JARs as shown below: ./WEB-INF/lib/aopalliance-1.0.jar./WEB-INF/lib/aspectjrt-1.6.10.jar./WEB-INF/lib/commons-fileupload-1.2.2.jar./WEB-INF/lib/commons-io-2.0.1.jar./WEB-INF/lib/el-api-2.2.jar./WEB-INF/lib/hibernate-validator-4.1.0.Final.jar./WEB-INF/lib/jackson-core-asl-1.8.1.jar./WEB-INF/lib/jackson-mapper-asl-1.8.1.jar./WEB-INF/lib/javax.inject-1.jar./WEB-INF/lib/jcl-over-slf4j-1.6.1.jar./WEB-INF/lib/jdom-1.0.jar./WEB-INF/lib/joda-time-1.6.2.jar./WEB-INF/lib/jstl-api-1.2.jar./WEB-INF/lib/jstl-impl-1.2.jar./WEB-INF/lib/log4j-1.2.16.jar./WEB-INF/lib/rome-1.0.0.jar./WEB-INF/lib/slf4j-api-1.6.1.jar./WEB-INF/lib/slf4j-log4j12-1.6.1.jar./WEB-INF/lib/spring-aop-3.1.0.RELEASE.jar./WEB-INF/lib/spring-asm-3.1.0.RELEASE.jar./WEB-INF/lib/spring-beans-3.1.0.RELEASE.jar./WEB-INF/lib/spring-context-3.1.0.RELEASE.jar./WEB-INF/lib/spring-context-support-3.1.0.RELEASE.jar./WEB-INF/lib/spring-core-3.1.0.RELEASE.jar./WEB-INF/lib/spring-expression-3.1.0.RELEASE.jar./WEB-INF/lib/spring-web-3.1.0.RELEASE.jar./WEB-INF/lib/spring-webmvc-3.1.0.RELEASE.jar./WEB-INF/lib/validation-api-1.0.0.GA.jar The app at least deployed and showed results this time. But still no database! Next I tried building "jpetstore" and got the error: [ERROR] Failed to execute goal on project org.springframework.samples.jpetstore:Could not resolve dependencies for project org.springframework.samples:org.springframework.samples.jpetstore:war:1.0.0-SNAPSHOT: Failed to collect dependencies for [commons-fileupload:commons-fileupload:jar:1.2.1 (compile), org.apache.struts:com.springsource.org.apache.struts:jar:1.2.9 (compile), javax.xml.rpc:com.springsource.javax.xml.rpc:jar:1.1.0 (compile), org.apache.commons:com.springsource.org.apache.commons.dbcp:jar:1.2.2.osgi (compile), commons-io:commons-io:jar:1.3.2 (compile), hsqldb:hsqldb:jar:1.8.0.7 (compile), org.apache.tiles:tiles-core:jar:2.2.0 (compile), org.apache.tiles:tiles-jsp:jar:2.2.0 (compile), org.tuckey:urlrewritefilter:jar:3.1.0 (compile), org.springframework:spring-webmvc:jar:3.0.0.BUILD-SNAPSHOT (compile), org.springframework:spring-orm:jar:3.0.0.BUILD-SNAPSHOT (compile), org.springframework:spring-context-support:jar:3.0.0.BUILD-SNAPSHOT (compile), org.springframework.webflow:spring-js:jar:2.0.7.RELEASE (compile), org.apache.ibatis:com.springsource.com.ibatis:jar:2.3.4.726 (runtime), com.caucho:com.springsource.com.caucho:jar:3.2.1 (compile), org.apache.axis:com.springsource.org.apache.axis:jar:1.4.0 (compile), javax.wsdl:com.springsource.javax.wsdl:jar:1.6.1 (compile), javax.servlet:jstl:jar:1.2 (runtime), org.aspectj:aspectjweaver:jar:1.6.5 (compile), javax.servlet:servlet-api:jar:2.5 (provided), javax.servlet.jsp:jsp-api:jar:2.1 (provided), junit:junit:jar:4.6 (test)]: Failed to read artifact descriptor for org.springframework:spring-webmvc:jar:3.0.0.BUILD-SNAPSHOT: Could not transfer artifact org.springframework:spring-webmvc:pom:3.0.0.BUILD-SNAPSHOT from/to JBoss repository (http://repository.jboss.com/maven2): Access denied to: http://repository.jboss.com/maven2/org/springframework/spring-webmvc/3.0.0.BUILD-SNAPSHOT/spring-webmvc-3.0.0.BUILD-SNAPSHOT.pom It appears the sample is broken - maybe I was pulling from the wrong repository - would be great if someone were to point me at a good target to use here. With a 50% hit on samples in this repository, I started searching through numerous blogs, most of which have either outdated information (using XML-heavy Spring 2.5), some piece of configuration (which is a typical "feature" of Spring) is missing, or too much complexity in the sample. I finally found this blog that worked like a charm. This blog creates a trivial Spring MVC 3 application using Hibernate and MySQL. This application performs CRUD operations on a single table in a database using typical Spring technologies.  I downloaded the sample code from the blog, deployed it on GlassFish 3.1.2 and could CRUD the "person" entity. The source code for this application can be downloaded here. More details on the application statistics below. And then I built a similar CRUD application in Java EE 6 using NetBeans wizards in a couple of minutes. The source code for the application can be downloaded here and the WAR here. The Spring Source Tool Suite may also offer similar wizard-driven capabilities but this blog focus primarily on comparing the runtimes. The lack of STS tutorials was slightly disappointing as well. NetBeans however has tons of text-based and video tutorials and tons of material even by the community. One more bit on the download size of tools bundle ... NetBeans 7.1.1 "All" is 211 MB (which includes GlassFish and Tomcat) Spring Tool Suite  2.9.0 is 347 MB (~ 65% bigger) This blog is not about the tooling comparison so back to the Java EE 6 version of the application .... In order to run the Java EE version on GlassFish, copy the MySQL Connector/J to glassfish3/glassfish/domains/domain1/lib/ext directory and create a JDBC connection pool and JDBC resource as: ./bin/asadmin create-jdbc-connection-pool --datasourceclassname \\ com.mysql.jdbc.jdbc2.optional.MysqlDataSource --restype \\ javax.sql.DataSource --property \\ portNumber=3306:user=mysql:password=mysql:databaseName=mydatabase \\ myConnectionPool ./bin/asadmin create-jdbc-resource --connectionpoolid myConnectionPool jdbc/myDataSource I generated WARs for the two projects and the table below highlights some differences between them: Java EE 6 Spring WAR File Size 0.021030 MB 10.87 MB (~516x) Number of files 20 53 (> 2.5x) Bundled libraries 0 36 Total size of libraries 0 12.1 MB XML files 3 5 LoC in XML files 50 (11 + 15 + 24) 129 (27 + 46 + 16 + 11 + 19) (~ 2.5x) Total .properties files 1 Bundle.properties 2 spring.properties, log4j.properties Cold Deploy 5,339 ms 11,724 ms Second Deploy 481 ms 6,261 ms Third Deploy 528 ms 5,484 ms Fourth Deploy 484 ms 5,576 ms Runtime memory ~73 MB ~101 MB Some points worth highlighting from the table ... 516x WAR file, 10x deployment time - With 12.1 MB of libraries (for a very basic application) bundled in your application, the WAR file size and the deployment time will naturally go higher. The WAR file for Spring-based application is 516x bigger and the deployment time is double during the first deployment and ~ 10x during subsequent deployments. The Java EE 6 application is fully portable and will run on any Java EE 6 compliant application server. 36 libraries in the WAR - There are 14 Java EE 6 compliant application servers today. Each of those servers provide all the functionality like transactions, dependency injection, security, persistence, etc typically required of an enterprise or web application. There is no need to bundle 36 libraries worth 12.1 MB for a trivial CRUD application. These 14 compliant application servers provide all the functionality baked in. Now you can also deploy these libraries in the container but then you don't get the "portability" offered by Spring in that case. Does your typical Spring deployment actually do that ? 3x LoC in XML - The number of XML files is about 1.6x and the LoC is ~ 2.5x. So much XML seems circa 2003 when the Java language had no annotations. The XML files can be further reduced, e.g. faces-config.xml can be replaced without providing i18n, but I just want to compare stock applications. Memory usage - Both the applications were deployed on default GlassFish 3.1.2 installation and any additional memory consumed as part of deployment/access was attributed to the application. This is by no means scientific but at least provides an initial ballpark. This area definitely needs more investigation. Another table that compares typical Java EE 6 compliant application servers and the custom-stack created for a Spring application ... Java EE 6 Spring Web Container ? 53 MB (tcServer 2.6.3 Developer Edition) Security ? 12 MB (Spring Security 3.1.0) Persistence ? 6.3 MB (Hibernate 4.1.0, required) Dependency Injection ? 5.3 MB (Framework) Web Services ? 796 KB (Spring WS 2.0.4) Messaging ? 3.4 MB (RabbitMQ Server 2.7.1) 936 KB (Java client 936) OSGi ? 1.3 MB (Spring OSGi 1.2.1) GlassFish and WebLogic (starting at 33 MB) 83.3 MB There are differentiating factors on both the stacks. But most of the functionality like security, persistence, and dependency injection is baked in a Java EE 6 compliant application server but needs to be individually managed and patched for a Spring application. This very quickly leads to a "stack explosion". The Java EE 6 servers are tested extensively on a variety of platforms in different combinations whereas a Spring application developer is responsible for testing with different JDKs, Operating Systems, Versions, Patches, etc. Oracle has both the leading OSS lightweight server with GlassFish and the leading enterprise Java server with WebLogic Server, both Java EE 6 and both with lightweight deployment options. The Web Container offered as part of a Java EE 6 application server not only deploys your enterprise Java applications but also provide operational management, diagnostics, and mission-critical capabilities required by your applications. The Java EE 6 platform also introduced the Web Profile which is a subset of the specifications from the entire platform. It is targeted at developers of modern web applications offering a reasonably complete stack, composed of standard APIs, and is capable out-of-the-box of addressing the needs of a large class of Web applications. As your applications grow, the stack can grow to the full Java EE 6 platform. The GlassFish Server Web Profile starting at 33MB (smaller than just the non-standard tcServer) provides most of the functionality typically required by a web application. WebLogic provides battle-tested functionality for a high throughput, low latency, and enterprise grade web application. No individual managing or patching, all tested and commercially supported for you! Note that VMWare does have a server, tcServer, but it is non-standard and not even certified to the level of the standard Web Profile most customers expect these days. Customers who choose this risk proprietary lock-in since VMWare does not seem to want to formally certify with either Java EE 6 Enterprise Platform or with Java EE 6 Web Profile but of course it would be great if they were to join the community and help their customers reduce the risk of deploying on VMWare software. Some more points to help you decide choose between Java EE 6 and Spring ... Freedom to choose container - There are 14 Java EE 6 compliant application servers today, with a variety of open source and commercial offerings. A Java EE 6 application can be deployed on any of those containers. So if you deployed your application on GlassFish today and would like to scale up with your demands then you can deploy the same application to WebLogic. And because of the portability of a Java EE 6 application, you can even take it a different vendor altogether. Spring requires a runtime which could be any of these app servers as well. But why use Spring when all the required functionality is already baked into the application server itself ? Spring also has a different definition of portability where they claim to bundle all the libraries in the WAR file and move to any application server. But we saw earlier how bloated that archive could be. The equivalent features in Spring runtime offerings (mainly tcServer) are not all open source, not as mature, and often require manual assembly.  Vendor choice - The Java EE 6 platform is created using the Java Community Process where all the big players like Oracle, IBM, RedHat, and Apache are conritbuting to make the platform successful. Each application server provides the basic Java EE 6 platform compliance and has its own competitive offerings. This allows you to choose an application server for deploying your Java EE 6 applications. If you are not happy with the support or feature of one vendor then you can move your application to a different vendor because of the portability promise offered by the platform. Spring is a set of products from a single company, one price book, one support organization, one sustaining organization, one sales organization, etc. If any of those cause a customer headache, where do you go ? Java EE, backed by multiple vendors, is a safer bet for those that are risk averse. Production support - With Spring, typically you need to get support from two vendors - VMWare and the container provider. With Java EE 6, all of this is typically provided by one vendor. For example, Oracle offers commercial support from systems, operating systems, JDK, application server, and applications on top of them. VMWare certainly offers complete production support but do you really want to put all your eggs in one basket ? Do you really use tcServer ? ;-) Maintainability - With Spring, you are likely building your own distribution with multiple JAR files, integrating, patching, versioning, etc of all those components. Spring's claim is that multiple JAR files allow you to go à la carte and pick the latest versions of different components. But who is responsible for testing whether all these versions work together ? Yep, you got it, its YOU! If something does not work, who patches and maintains the JARs ? Of course, you! Commercial support for such a configuration ? On your own! The Java EE application servers manage all of this for you and provide a well-tested and commercially supported bundle. While it is always good to realize that there is something new and improved that updates and replaces older frameworks like Spring, the good news is not only does a Java EE 6 container offer what is described here, most also will let you deploy and run your Spring applications on them while you go through an upgrade to a more modern architecture. End result, you get the best of both worlds - keeping your legacy investment but moving to a more agile, lightweight world of Java EE 6. A message to the Spring lovers ... The complexity in J2EE 1.2, 1.3, and 1.4 led to the genesis of Spring but that was in 2004. This is 2012 and the name has changed to "Java EE 6" :-) There are tons of improvements in the Java EE platform to make it easy-to-use and powerful. Some examples: Adding @Stateless on a POJO makes it an EJB EJBs can be packaged in a WAR with no special packaging or deployment descriptors "web.xml" and "faces-config.xml" are optional in most of the common cases Typesafe dependency injection is now part of the Java EE platform Add @Path on a POJO allows you to publish it as a RESTful resource EJBs can be used as backing beans for Facelets-driven JSF pages providing full MVC Java EE 6 WARs are known to be kilobytes in size and deployed in milliseconds Tons of other simplifications in the platform and application servers So if you moved away from J2EE to Spring many years ago and have not looked at Java EE 6 (which has been out since Dec 2009) then you should definitely try it out. Just be at least aware of what other alternatives are available instead of restricting yourself to one stack. Here are some workshops and screencasts worth trying: screencast #37 shows how to build an end-to-end application using NetBeans screencast #36 builds the same application using Eclipse javaee-lab-feb2012.pdf is a 3-4 hours self-paced hands-on workshop that guides you to build a comprehensive Java EE 6 application using NetBeans Each city generally has a "spring cleanup" program every year. It allows you to clean up the mess from your house. For your software projects, you don't need to wait for an annual event, just get started and reduce the technical debt now! Move away from your legacy Spring-based applications to a lighter and more modern approach of building enterprise Java applications using Java EE 6. Watch this beautiful presentation that explains how to migrate from Spring -> Java EE 6: List of files in the Java EE 6 project: ./index.xhtml./META-INF./person./person/Create.xhtml./person/Edit.xhtml./person/List.xhtml./person/View.xhtml./resources./resources/css./resources/css/jsfcrud.css./template.xhtml./WEB-INF./WEB-INF/classes./WEB-INF/classes/Bundle.properties./WEB-INF/classes/META-INF./WEB-INF/classes/META-INF/persistence.xml./WEB-INF/classes/org./WEB-INF/classes/org/javaee./WEB-INF/classes/org/javaee/javaeemysql./WEB-INF/classes/org/javaee/javaeemysql/AbstractFacade.class./WEB-INF/classes/org/javaee/javaeemysql/Person.class./WEB-INF/classes/org/javaee/javaeemysql/Person_.class./WEB-INF/classes/org/javaee/javaeemysql/PersonController$1.class./WEB-INF/classes/org/javaee/javaeemysql/PersonController$PersonControllerConverter.class./WEB-INF/classes/org/javaee/javaeemysql/PersonController.class./WEB-INF/classes/org/javaee/javaeemysql/PersonFacade.class./WEB-INF/classes/org/javaee/javaeemysql/util./WEB-INF/classes/org/javaee/javaeemysql/util/JsfUtil.class./WEB-INF/classes/org/javaee/javaeemysql/util/PaginationHelper.class./WEB-INF/faces-config.xml./WEB-INF/web.xml List of files in the Spring 3.x project: ./META-INF ./META-INF/MANIFEST.MF./WEB-INF./WEB-INF/applicationContext.xml./WEB-INF/classes./WEB-INF/classes/log4j.properties./WEB-INF/classes/org./WEB-INF/classes/org/krams ./WEB-INF/classes/org/krams/tutorial ./WEB-INF/classes/org/krams/tutorial/controller ./WEB-INF/classes/org/krams/tutorial/controller/MainController.class ./WEB-INF/classes/org/krams/tutorial/domain ./WEB-INF/classes/org/krams/tutorial/domain/Person.class ./WEB-INF/classes/org/krams/tutorial/service ./WEB-INF/classes/org/krams/tutorial/service/PersonService.class ./WEB-INF/hibernate-context.xml ./WEB-INF/hibernate.cfg.xml ./WEB-INF/jsp ./WEB-INF/jsp/addedpage.jsp ./WEB-INF/jsp/addpage.jsp ./WEB-INF/jsp/deletedpage.jsp ./WEB-INF/jsp/editedpage.jsp ./WEB-INF/jsp/editpage.jsp ./WEB-INF/jsp/personspage.jsp ./WEB-INF/lib ./WEB-INF/lib/antlr-2.7.6.jar ./WEB-INF/lib/aopalliance-1.0.jar ./WEB-INF/lib/c3p0-0.9.1.2.jar ./WEB-INF/lib/cglib-nodep-2.2.jar ./WEB-INF/lib/commons-beanutils-1.8.3.jar ./WEB-INF/lib/commons-collections-3.2.1.jar ./WEB-INF/lib/commons-digester-2.1.jar ./WEB-INF/lib/commons-logging-1.1.1.jar ./WEB-INF/lib/dom4j-1.6.1.jar ./WEB-INF/lib/ejb3-persistence-1.0.2.GA.jar ./WEB-INF/lib/hibernate-annotations-3.4.0.GA.jar ./WEB-INF/lib/hibernate-commons-annotations-3.1.0.GA.jar ./WEB-INF/lib/hibernate-core-3.3.2.GA.jar ./WEB-INF/lib/javassist-3.7.ga.jar ./WEB-INF/lib/jstl-1.1.2.jar ./WEB-INF/lib/jta-1.1.jar ./WEB-INF/lib/junit-4.8.1.jar ./WEB-INF/lib/log4j-1.2.14.jar ./WEB-INF/lib/mysql-connector-java-5.1.14.jar ./WEB-INF/lib/persistence-api-1.0.jar ./WEB-INF/lib/slf4j-api-1.6.1.jar ./WEB-INF/lib/slf4j-log4j12-1.6.1.jar ./WEB-INF/lib/spring-aop-3.0.5.RELEASE.jar ./WEB-INF/lib/spring-asm-3.0.5.RELEASE.jar ./WEB-INF/lib/spring-beans-3.0.5.RELEASE.jar ./WEB-INF/lib/spring-context-3.0.5.RELEASE.jar ./WEB-INF/lib/spring-context-support-3.0.5.RELEASE.jar ./WEB-INF/lib/spring-core-3.0.5.RELEASE.jar ./WEB-INF/lib/spring-expression-3.0.5.RELEASE.jar ./WEB-INF/lib/spring-jdbc-3.0.5.RELEASE.jar ./WEB-INF/lib/spring-orm-3.0.5.RELEASE.jar ./WEB-INF/lib/spring-tx-3.0.5.RELEASE.jar ./WEB-INF/lib/spring-web-3.0.5.RELEASE.jar ./WEB-INF/lib/spring-webmvc-3.0.5.RELEASE.jar ./WEB-INF/lib/standard-1.1.2.jar ./WEB-INF/lib/xml-apis-1.0.b2.jar ./WEB-INF/spring-servlet.xml ./WEB-INF/spring.properties ./WEB-INF/web.xml So, are you excited about Java EE 6 ? Want to get started now ? Here are some resources: Java EE 6 SDK (including runtime, samples, tutorials etc) GlassFish Server Open Source Edition 3.1.2 (Community) Oracle GlassFish Server 3.1.2 (Commercial) Java EE 6 using WebLogic 12c and NetBeans (Video) Java EE 6 with NetBeans and GlassFish (Video) Java EE with Eclipse and GlassFish (Video)

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