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• Abstracting functionality

  - by Ralf Westphal

  Originally posted on: http://geekswithblogs.net/theArchitectsNapkin/archive/2014/08/22/abstracting-functionality.aspxWhat is more important than data? Functionality. Yes, I strongly believe we should switch to a functionality over data mindset in programming. Or actually switch back to it. Focus on functionality Functionality once was at the core of software development. Back when algorithms were the first thing you heard about in CS classes. Sure, data structures, too, were important - but always from the point of view of algorithms. (Niklaus Wirth gave one of his books the title “Algorithms + Data Structures” instead of “Data Structures + Algorithms” for a reason.) The reason for the focus on functionality? Firstly, because software was and is about doing stuff. Secondly because sufficient performance was hard to achieve, and only thirdly memory efficiency. But then hardware became more powerful. That gave rise to a new mindset: object orientation. And with it functionality was devalued. Data took over its place as the most important aspect. Now discussions revolved around structures motivated by data relationships. (John Beidler gave his book the title “Data Structures and Algorithms: An Object Oriented Approach” instead of the other way around for a reason.) Sure, this data could be embellished with functionality. But nevertheless functionality was second. When you look at (domain) object models what you mostly find is (domain) data object models. The common object oriented approach is: data aka structure over functionality. This is true even for the most modern modeling approaches like Domain Driven Design. Look at the literature and what you find is recommendations on how to get data structures right: aggregates, entities, value objects. I´m not saying this is what object orientation was invented for. But I´m saying that´s what I happen to see across many teams now some 25 years after object orientation became mainstream through C++, Delphi, and Java. But why should we switch back? Because software development cannot become truly agile with a data focus. The reason for that lies in what customers need first: functionality, behavior, operations. To be clear, that´s not why software is built. The purpose of software is to be more efficient than the alternative. Money mainly is spent to get a certain level of quality (e.g. performance, scalability, security etc.). But without functionality being present, there is nothing to work on the quality of. What customers want is functionality of a certain quality. ASAP. And tomorrow new functionality needs to be added, existing functionality needs to be changed, and quality needs to be increased. No customer ever wanted data or structures. Of course data should be processed. Data is there, data gets generated, transformed, stored. But how the data is structured for this to happen efficiently is of no concern to the customer. Ask a customer (or user) whether she likes the data structured this way or that way. She´ll say, “I don´t care.” But ask a customer (or user) whether he likes the functionality and its quality this way or that way. He´ll say, “I like it” (or “I don´t like it”). Build software incrementally From this very natural focus of customers and users on functionality and its quality follows we should develop software incrementally. That´s what Agility is about. Deliver small increments quickly and often to get frequent feedback. That way less waste is produced, and learning can take place much easier (on the side of the customer as well as on the side of developers). An increment is some added functionality or quality of functionality.[1] So as it turns out, Agility is about functionality over whatever. But software developers’ thinking is still stuck in the object oriented mindset of whatever over functionality. Bummer. I guess that (at least partly) explains why Agility always hits a glass ceiling in projects. It´s a clash of mindsets, of cultures. Driving software development by demanding small increases in functionality runs against thinking about software as growing (data) structures sprinkled with functionality. (Excuse me, if this sounds a bit broad-brush. But you get my point.) The need for abstraction In the end there need to be data structures. Of course. Small and large ones. The phrase functionality over data does not deny that. It´s not functionality instead of data or something. It´s just over, i.e. functionality should be thought of first. It´s a tad more important. It´s what the customer wants. That´s why we need a way to design functionality. Small and large. We need to be able to think about functionality before implementing it. We need to be able to reason about it among team members. We need to be able to communicate our mental models of functionality not just by speaking about them, but also on paper. Otherwise reasoning about it does not scale. We learned thinking about functionality in the small using flow charts, Nassi-Shneiderman diagrams, pseudo code, or UML sequence diagrams. That´s nice and well. But it does not scale. You can use these tools to describe manageable algorithms. But it does not work for the functionality triggered by pressing the “1-Click Order” on an amazon product page for example. There are several reasons for that, I´d say. Firstly, the level of abstraction over code is negligible. It´s essentially non-existent. Drawing a flow chart or writing pseudo code or writing actual code is very, very much alike. All these tools are about control flow like code is.[2] In addition all tools are computationally complete. They are about logic which is expressions and especially control statements. Whatever you code in Java you can fully (!) describe using a flow chart. And then there is no data. They are about control flow and leave out the data altogether. Thus data mostly is assumed to be global. That´s shooting yourself in the foot, as I hope you agree. Even if it´s functionality over data that does not mean “don´t think about data”. Right to the contrary! Functionality only makes sense with regard to data. So data needs to be in the picture right from the start - but it must not dominate the thinking. The above tools fail on this. Bottom line: So far we´re unable to reason in a scalable and abstract manner about functionality. That´s why programmers are so driven to start coding once they are presented with a problem. Programming languages are the only tool they´ve learned to use to reason about functional solutions. Or, well, there might be exceptions. Mathematical notation and SQL may have come to your mind already. Indeed they are tools on a higher level of abstraction than flow charts etc. That´s because they are declarative and not computationally complete. They leave out details - in order to deliver higher efficiency in devising overall solutions. We can easily reason about functionality using mathematics and SQL. That´s great. Except for that they are domain specific languages. They are not general purpose. (And they don´t scale either, I´d say.) Bummer. So to be more precise we need a scalable general purpose tool on a higher than code level of abstraction not neglecting data. Enter: Flow Design. Abstracting functionality using data flows I believe the solution to the problem of abstracting functionality lies in switching from control flow to data flow. Data flow very naturally is not about logic details anymore. There are no expressions and no control statements anymore. There are not even statements anymore. Data flow is declarative by nature. With data flow we get rid of all the limiting traits of former approaches to modeling functionality. In addition, nomen est omen, data flows include data in the functionality picture. With data flows, data is visibly flowing from processing step to processing step. Control is not flowing. Control is wherever it´s needed to process data coming in. That´s a crucial difference and needs some rewiring in your head to be fully appreciated.[2] Since data flows are declarative they are not the right tool to describe algorithms, though, I´d say. With them you don´t design functionality on a low level. During design data flow processing steps are black boxes. They get fleshed out during coding. Data flow design thus is more coarse grained than flow chart design. It starts on a higher level of abstraction - but then is not limited. By nesting data flows indefinitely you can design functionality of any size, without losing sight of your data. Data flows scale very well during design. They can be used on any level of granularity. And they can easily be depicted. Communicating designs using data flows is easy and scales well, too. The result of functional design using data flows is not algorithms (too low level), but processes. Think of data flows as descriptions of industrial production lines. Data as material runs through a number of processing steps to be analyzed, enhances, transformed. On the top level of a data flow design might be just one processing step, e.g. “execute 1-click order”. But below that are arbitrary levels of flows with smaller and smaller steps. That´s not layering as in “layered architecture”, though. Rather it´s a stratified design à la Abelson/Sussman. Refining data flows is not your grandpa´s functional decomposition. That was rooted in control flows. Refining data flows does not suffer from the limits of functional decomposition against which object orientation was supposed to be an antidote. Summary I´ve been working exclusively with data flows for functional design for the past 4 years. It has changed my life as a programmer. What once was difficult is now easy. And, no, I´m not using Clojure or F#. And I´m not a async/parallel execution buff. Designing the functionality of increments using data flows works great with teams. It produces design documentation which can easily be translated into code - in which then the smallest data flow processing steps have to be fleshed out - which is comparatively easy. Using a systematic translation approach code can mirror the data flow design. That way later on the design can easily be reproduced from the code if need be. And finally, data flow designs play well with object orientation. They are a great starting point for class design. But that´s a story for another day. To me data flow design simply is one of the missing links of systematic lightweight software design. There are also other artifacts software development can produce to get feedback, e.g. process descriptions, test cases. But customers can be delighted more easily with code based increments in functionality. ? No, I´m not talking about the endless possibilities this opens for parallel processing. Data flows are useful independently of multi-core processors and Actor-based designs. That´s my whole point here. Data flows are good for reasoning and evolvability. So forget about any special frameworks you might need to reap benefits from data flows. None are necessary. Translating data flow designs even into plain of Java is possible. ?

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• PASS: Bylaw Changes

  - by Bill Graziano

  While you’re reading this, a post should be going up on the PASS blog on the plans to change our bylaws.  You should be able to find our old bylaws, our proposed bylaws and a red-lined version of the changes.  We plan to listen to feedback until March 31st.  At that point we’ll decide whether to vote on these changes or take other action. The executive summary is that we’re adding a restriction to prevent more than two people from the same company on the Board and eliminating the Board’s Officer Appointment Committee to have Officers directly elected by the Board.  This second change better matches how officer elections have been conducted in the past. The Gritty Details Our scope was to change bylaws to match how PASS actually works and tackle a limited set of issues.  Changing the bylaws is hard.  We’ve been working on these changes since the March board meeting last year.  At that meeting we met and talked through the issues we wanted to address.  In years past the Board has tried to come up with language and then we’ve discussed and negotiated to get to the result.  In March, we gave HQ guidance on what we wanted and asked them to come up with a starting point.  Hannes worked on building us an initial set of changes that we could work our way through.  Discussing changes like this over email is difficult wasn’t very productive.  We do a much better job on this at the in-person Board meetings.  Unfortunately there are only 2 or 3 of those a year. In August we met in Nashville and spent time discussing the changes.  That was also the day after we released the slate for the 2010 election. The discussion around that colored what we talked about in terms of these changes.  We talked very briefly at the Summit and again reviewed and revised the changes at the Board meeting in January.  This is the result of those changes and discussions. We made numerous small changes to clean up language and make wording more clear.  We also made two big changes. Director Employment Restrictions The first is that only two people from the same company can serve on the Board at the same time.  The actual language in section VI.3 reads: A maximum of two (2) Directors who are employed by, or who are joint owners or partners in, the same for-profit venture, company, organization, or other legal entity, may concurrently serve on the PASS Board of Directors at any time. The definition of “employed” is at the sole discretion of the Board. And what a mess this turns out to be in practice.  Our membership is a hodgepodge of interlocking relationships.  Let’s say three Board members get together and start a blog service for SQL Server bloggers.  It’s technically for-profit.  Let’s assume it makes $8 in the first year.  Does that trigger this clause?  (Technically yes.)  We had a horrible time trying to write language that covered everything.  All the sample bylaws that we found were just as vague as this. That led to the third clause in this section.  The first sentence reads: The Board of Directors reserves the right, strictly on a case-by-case basis, to overrule the requirements of Section VI.3 by majority decision for any single Director’s conflict of employment. We needed some way to handle the trivial issues and exercise some judgment.  It seems like a public vote is the best way.  This discloses the relationship and gets each Board member on record on the issue.   In practice I think this clause will rarely be used.  I think this entire section will only be invoked for actual employment issues and not for small side projects.  In either case we have the mechanisms in place to handle it in a public, transparent way. That’s the first and third clauses.  The second clause says that if your situation changes and you fall afoul of this restriction you need to notify the Board.  The clause further states that if this new job means a Board members violates the “two-per-company” rule the Board may request their resignation.  The Board can also  allow the person to continue serving with a majority vote.  I think this will also take some judgment.  Consider a person switching jobs that leads to three people from the same company.  I’m very likely to ask for someone to resign if all three are two weeks into a two year term.  I’m unlikely to ask anyone to resign if one is two weeks away from ending their term.  In either case, the decision will be a public vote that we can be held accountable for. One concern that was raised was whether this would affect someone choosing to accept a job.  I think that’s a choice for them to make.  PASS is clearly stating its intent that only two directors from any one organization should serve at any time.  Once these bylaws are approved, this policy should not come as a surprise to any potential or current Board members considering a job change.  This clause isn’t perfect.  The biggest hole is business relationships that aren’t defined above.  Let’s say that two employees from company “X” serve on the Board.  What happens if I accept a full-time consulting contract with that company?  Let’s assume I’m working directly for one of the two existing Board members.  That doesn’t violate section VI.3.  But I think it’s clearly the kind of relationship we’d like to prevent.  Unfortunately that was even harder to write than what we have now.  I fully expect that in the next revision of the bylaws we’ll address this.  It just didn’t make it into this one. Officer Elections The officer election process received a slightly different rewrite.  Our goal was to codify in the bylaws the actual process we used to elect the officers.  The officers are the President, Executive Vice-President (EVP) and Vice-President of Marketing.  The Immediate Past President (IPP) is also an officer but isn’t elected.  The IPP serves in that role for two years after completing their term as President.  We do that for continuity’s sake.  Some organizations have a President-elect that serves for one or two years.  The group that founded PASS chose to have an IPP. When I started on the Board, the Nominating Committee (NomCom) selected the slate for the at-large directors and the slate for the officers.  There was always one candidate for each officer position.  It wasn’t really an election so much as the NomCom decided who the next person would be for each officer position.  Behind the scenes the Board worked to select the best people for the role. In June 2009 that process was changed to bring it line with what actually happens.  An Officer Appointment Committee was created that was a subset of the Board.  That committee would take time to interview the candidates and present a slate to the Board for approval.  The majority vote of the Board would determine the officers for the next two years.  In practice the Board itself interviewed the candidates and conducted the elections.  That means it was time to change the bylaws again. Section VII.2 and VII.3 spell out the process used to select the officers.  We use the phrase “Officer Appointment” to separate it from the Director election but the end result is that the Board elects the officers.  Section VII.3 starts: Officers shall be appointed bi-annually by a majority of all the voting members of the Board of Directors. Everything else revolves around that sentence.  We use the word appoint but they truly are elected.  There are details in the bylaws for term limits, minimum requirements for President (1 prior term as an officer), tie breakers and filling vacancies. In practice we will have an election for President, then an election for EVP and then an election for VP Marketing.  That means that losing candidates will be able to fall down the ladder and run for the next open position.  Another point to note is that officers aren’t at-large directors.  That means if a current sitting officer loses all three elections they are off the Board.  Having Board member votes public will help with the transparency of this approach. This process has a number of positive and negatives.  The biggest concern I expect to hear is that our members don’t directly choose the officers.  I’m going to try and list all the positives and negatives of this approach. Many non-profits value continuity and are slower to change than a business.  On the plus side this promotes that.  On the negative side this promotes that.  If we change too slowly the members complain that we aren’t responsive.  If we change too quickly we make mistakes and fail at various things.  We’ve been criticized for both of those lately so I’m not entirely sure where to draw the line.  My rough assumption to this point is that we’re going too slow on governance and too quickly on becoming “more than a Summit.”  This approach creates competition in the officer elections.  If you are an at-large director there is no consequence to losing an election.  If you are an officer the only way to stay on the Board is to win an officer election or an at-large election.  If you are an officer and lose an election you can always run for the next office down.  This makes it very easy for multiple people to contest an election. There is value in a person moving through the officer positions up to the Presidency.  Having the Board select the officers promotes this.  The down side is that it takes a LOT of time to get to the Presidency.  We’ve had good people struggle with burnout.  We’ve had lots of discussion around this.  The process as we’ve described it here makes it possible for someone to move quickly through the ranks but doesn’t prevent people from working their way up through each role. We talked long and hard about having the officers elected by the members.  We had a self-imposed deadline to complete these changes prior to elections this summer. The other challenge was that our original goal was to make the bylaws reflect our actual process rather than create a new one.  I believe we accomplished this goal. We ran out of time to consider this option in the detail it needs.  Having member elections for officers needs a number of problems solved.  We would need a way for candidates to fall through the election.  This is what promotes competition.  Without this few people would risk an election and we’ll be back to one candidate per slot.  We need to do this without having multiple elections.  We may be able to copy what other organizations are doing but I was surprised at how little I could find on other organizations.  We also need a way for people that lose an officer election to win an at-large election.  Otherwise we’ll have very little competition for officers. This brings me to an area that I think we as a Board haven’t done a good job.  We haven’t built a strong process to tell you who is doing a good job and who isn’t.  This is a double-edged sword.  I don’t want to highlight Board members that are failing.  That’s not a good way to get people to volunteer and run for the Board.  But I also need a way let the members make an informed choice about who is doing a good job and would make a good officer.  Encouraging Board members to blog, publishing minutes and making votes public helps in that regard but isn’t the final answer.  I don’t know what the final answer is yet.  I do know that the Board members themselves are uniquely positioned to know which other Board members are doing good work.  They know who speaks up in meetings, who works to build consensus, who has good ideas and who works with the members.  What I Could Do Better I’ve learned a lot writing this about how we communicated with our members.  The next time we revise the bylaws I’d do a few things differently.  The biggest change would be to provide better documentation.  The March 2009 minutes provide a very detailed look into what changes we wanted to make to the bylaws.  Looking back, I’m a little surprised at how closely they matched our final changes and covered the various arguments.  If you just read those you’d get 90% of what we eventually changed.  Nearly everything else was just details around implementation.  I’d also consider publishing a scope document defining exactly what we were doing any why.  I think it really helped that we had a limited, defined goal in mind.  I don’t think we did a good job communicating that goal outside the meeting minutes though. That said, I wish I’d blogged more after the August and January meeting.  I think it would have helped more people to know that this change was coming and to be ready for it. Conclusion These changes address two big concerns that the Board had.  First, it prevents a single organization from dominating the Board.  Second, it codifies and clearly spells out how officers are elected.  This is the process that was previously followed but it was somewhat murky.  These changes bring clarity to this and clearly explain the process the Board will follow. We’re going to listen to feedback until March 31st.  At that time we’ll decide whether to approve these changes.  I’m also assuming that we’ll start another round of changes in the next year or two.  Are there other issues in the bylaws that we should tackle in the future?

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• PASS: Bylaw Change 2013

  - by Bill Graziano

  PASS launched a Global Growth Initiative in the Summer of 2011 with the appointment of three international Board advisors.  Since then we’ve thought and talked extensively about how we make PASS more relevant to our members outside the US and Canada.  We’ve collected much of that discussion in our Global Growth site.  You can find vision documents, plans, governance proposals, feedback sites, and transcripts of Twitter chats and town hall meetings.  We also address these plans at the Board Q&A during the 2012 Summit. One of the biggest changes coming out of this process is around how we elect Board members.  And that requires a change to the bylaws.  We published the proposed bylaw changes as a red-lined document so you can clearly see the changes.  Our goal in these bylaw changes was to address the changes required by the global growth initiatives, conduct a legal review of the document and address other minor issues in the document.  There are numerous small wording changes throughout the document.  For example, we replaced every reference of “The Corporation” with the word “PASS” so it now reads “PASS is organized…”. Board Composition The biggest change in these bylaw changes is how the Board is composed and elected.  This discussion starts in section VI.2.  This section now says that some elected directors will come from geographic regions.  I think this is the best way to make sure we give all of our members a voice in the leadership of the organization.  The key parts of this section are: The remaining Directors (i.e. the non-Officer Directors and non-Vendor Appointed Directors) shall be elected by the voting membership (“Elected Directors”). Elected Directors shall include representatives of defined PASS regions (“Regions”) as set forth below (“Regional Directors”) and at minimum one (1) additional Director-at-Large whose selection is not limited by region. Regional Directors shall include, but are not limited to, two (2) seats for the Region covering Canada and the United States of America. Additional Regions for the purpose of electing additional Regional Directors and additional Director-at-Large seats for the purpose of expanding the Board shall be defined by a majority vote of the current Board of Directors and must be established prior to the public call for nominations in the general election. Previously defined Regions and seats approved by the Board of Directors shall remain in effect and can only be modified by a 2/3 majority vote by the then current Board of Directors. Currently PASS has six At-Large Directors elected by the members.  These changes allow for a Regional Director position that is elected by the members but must come from a particular region.  It also stipulates that there must always be at least one Director-at-Large who can come from any region. We also understand that PASS is currently a very US-centric organization.  Our Summit is held in America, roughly half our chapters are in the US and Canada and most of the Board members over the last ten years have come from America.  We wanted to reflect that by making sure that our US and Canadian volunteers would continue to play a significant role by ensuring that two Regional seats are reserved specifically for Canada and the US. Other than that, the bylaws don’t create any specific regional seats.  These rules allow us to create Regional Director seats but don’t require it.  We haven’t fully discussed what the criteria will be in order for a region to have a seat designated for it or how many regions there will be.  In our discussions we’ve broadly discussed regions for United States and Canada Europe, Middle East, and Africa (EMEA) Australia, New Zealand and Asia (also known as Asia Pacific or APAC) Mexico, South America, and Central America (LATAM) As you can see, our thinking is that there will be a few large regions.  I’ve also considered a non-North America region that we can gradually split into the regions above as our membership grows in those areas.  The regions will be defined by a policy document that will be published prior to the elections. I’m hoping that over the next year we can begin to publish more of what we do as Board-approved policy documents. While the bylaws only require a single non-region specific At-large Director, I would expect we would always have two.  That way we can have one in each election.  I think it’s important that we always have one seat open that anyone who is eligible to run for the Board can contest.  The Board is required to have any regions defined prior to the start of the election process. Board Elections – Regional Seats We spent a lot of time discussing how the elections would work for these Regional Director seats.  Ultimately we decided that the simplest solution is that every PASS member should vote for every open seat.  Section VIII.3 reads: Candidates who are eligible (i.e. eligible to serve in such capacity subject to the criteria set forth herein or adopted by the Board of Directors) shall be designated to fill open Board seats in the following order of priority on the basis of total votes received: (i) full term Regional Director seats, (ii) full term Director-at-Large seats, (iii) not full term (vacated) Regional Director seats, (iv) not full term (vacated) Director-at-Large seats. For the purposes of clarity, because of eligibility requirements, it is contemplated that the candidates designated to the open Board seats may not receive more votes than certain other candidates who are not selected to the Board. We debated whether to have multiple ballots or one single ballot.  Multiple ballot elections get complicated quickly.  Let’s say we have a ballot for US/Canada and one for Region 2.  After that we’d need a mechanism to merge those two together and come up with the winner of the at-large seat or have another election for the at-large position.  We think the best way to do this is a single ballot and putting the highest vote getters into the most restrictive seats.  Let’s look at an example: There are seats open for Region 1, Region 2 and at-large.  The election results are as follows: Candidate A (eligible for Region 1) – 550 votes Candidate B (eligible for Region 1) – 525 votes Candidate C (eligible for Region 1) – 475 votes Candidate D (eligible for Region 2) – 125 votes Candidate E (eligible for Region 2) – 75 votes In this case, Candidate A is the winner for Region 1 and is assigned that seat.  Candidate D is the winner for Region 2 and is assigned that seat.  The at-large seat is filled by the high remaining vote getter which is Candidate B. The key point to understand is that we may have a situation where a person with a lower vote total is elected to a regional seat and a person with a higher vote total is excluded.  This will be true whether we had multiple ballots or a single ballot.  Board Elections – Vacant Seats The other change to the election process is for vacant Board seats.  The actual changes are sprinkled throughout the document. Previously we didn’t have a mechanism that allowed for an election of a Board seat that we knew would be vacant in the future.  The most common case is when a Board members moves to an Officer role in the middle of their term.  One of the key changes is to allow the number of votes members have to match the number of open seats.  This allows each voter to express their preference on all open seats.  This only applies when we know about the opening prior to the call for nominations.  This all means that if there’s a seat will be open at the start of the next Board term, and we know about it prior to the call for nominations, we can include that seat in the elections.  Ultimately, the aim is to have PASS members decide who sits on the Board in as many situations as possible. We discussed the option of changing the bylaws to just take next highest vote-getter in all other cases.  I think that’s wrong for the following reasons: All voters aren’t able to express an opinion on all candidates.  If there are five people running for three seats, you can only vote for three.  You have no way to express your preference between #4 and #5. Different candidates may have different information about the number of seats available.  A person may learn that a Board member plans to resign at the end of the year prior to that information being made public. They may understand that the top four vote getters will end up on the Board while the rest of the members believe there are only three openings.  This may affect someone’s decision to run.  I don’t think this creates a transparent, fair election. Board members may use their knowledge of the election results to decide whether to remain on the Board or not.  Admittedly this one is unlikely but I don’t want to create a situation where this accusation can be leveled. I think the majority of vacancies in the future will be handled through elections.  The bylaw section quoted above also indicates that partial term vacancies will be filled after the full term seats are filled. Removing Directors Section VI.7 on removing directors has always had a clause that allowed members to remove an elected director.  We also had a clause that allowed appointed directors to be removed.  We added a clause that allows the Board to remove for cause any director with a 2/3 majority vote.  The updated text reads: Any Director may be removed for cause by a 2/3 majority vote of the Board of Directors whenever in its judgment the best interests of PASS would be served thereby. Notwithstanding the foregoing, the authority of any Director to act as in an official capacity as a Director or Officer of PASS may be suspended by the Board of Directors for cause. Cause for suspension or removal of a Director shall include but not be limited to failure to meet any Board-approved performance expectations or the presence of a reason for suspension or dismissal as listed in Addendum B of these Bylaws. The first paragraph is updated and the second and third are unchanged (except cleaning up language).  If you scroll down and look at Addendum B of these bylaws you find the following: Cause for suspension or dismissal of a member of the Board of Directors may include: Inability to attend Board meetings on a regular basis. Inability or unwillingness to act in a capacity designated by the Board of Directors. Failure to fulfill the responsibilities of the office. Inability to represent the Region elected to represent Failure to act in a manner consistent with PASS's Bylaws and/or policies. Misrepresentation of responsibility and/or authority. Misrepresentation of PASS. Unresolved conflict of interests with Board responsibilities. Breach of confidentiality. The bold line about your inability to represent your region is what we added to the bylaws in this revision.  We also added a clause to section VII.3 allowing the Board to remove an officer.  That clause is much less restrictive.  It doesn’t require cause and only requires a simple majority. The Board of Directors may remove any Officer whenever in their judgment the best interests of PASS shall be served by such removal. Other There are numerous other small changes throughout the document. Proxy voting.  The laws around how members and Board members proxy votes are specific in Illinois law.  PASS is an Illinois corporation and is subject to Illinois laws.  We changed section IV.5 to come into compliance with those laws.  Specifically this says you can only vote through a proxy if you have a written proxy through your authorized attorney.  English language proficiency.  As we increase our global footprint we come across more members that aren’t native English speakers.  The business of PASS is conducted in English and it’s important that our Board members speak English.  If we get big enough to afford translators, we may be able to relax this but right now we need English language skills for effective Board members. Committees.  The language around committees in section IX is old and dated.  Our lawyers advised us to clean it up.  This section specifically applies to any committees that the Board may form outside of portfolios.  We removed the term limits, quorum and vacancies clause.  We don’t currently have any committees that this would apply to.  The Nominating Committee is covered elsewhere in the bylaws. Electronic Votes.  The change allows the Board to vote via email but the results must be unanimous.  This is to conform with Illinois state law. Immediate Past President.  There was no mechanism to fill the IPP role if an outgoing President chose not to participate.  We changed section VII.8 to allow the Board to invite any previous President to fill the role by majority vote. Nominations Committee.  We’ve opened the language to allow for the transparent election of the Nominations Committee as outlined by the 2011 Election Review Committee. Revocation of Charters. The language surrounding the revocation of charters for local groups was flagged by the lawyers. We have allowed for the local user group to make all necessary payment before considering returning of items to PASS if required. Bylaw notification. We’ve spent countless meetings working on these bylaws with the intent to not open them again any time in the near future. Should the bylaws be opened again, we have included a clause ensuring that the PASS membership is involved. I’m proud that the Board has remained committed to transparency and accountability to members. This clause will require that same level of commitment in the future even when all the current Board members have rolled off. I think that covers everything.  I’d encourage you to look through the red-line document and see the changes.  It’s helpful to look at the language that’s being removed and the language that’s being added.  I’m happy to answer any questions here or you can email them to [email protected].

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• Tips for XNA WP7 Developers

  - by Michael B. McLaughlin

  There are several things any XNA developer should know/consider when coming to the Windows Phone 7 platform. This post assumes you are familiar with the XNA Framework and with the changes between XNA 3.1 and XNA 4.0. It’s not exhaustive; it’s simply a list of things I’ve gathered over time. I may come back and add to it over time, and I’m happy to add anything anyone else has experienced or learned as well. Display · The screen is either 800x480 or 480x800. · But you aren’t required to use only those resolutions. · The hardware scaler on the phone will scale up from 240x240. · One dimension will be capped at 800 and the other at 480; which depends on your code, but you cannot have, e.g., an 800x600 back buffer – that will be created as 800x480. · The hardware scaler will not normally change aspect ratio, though, so no unintended stretching. · Any dimension (width, height, or both) below 240 will be adjusted to 240 (without any aspect ratio adjustment such that, e.g. 200x240 will be treated as 240x240). · Dimensions below 240 will be honored in terms of calculating whether to use portrait or landscape. · If dimensions are exactly equal or if height is greater than width then game will be in portrait. · If width is greater than height, the game will be in landscape. · Landscape games will automatically flip if the user turns the phone 180°; no code required. · Default landscape is top = left. In other words a user holding a phone who starts a landscape game will see the first image presented so that the “top” of the screen is along the right edge of his/her phone, such that the natural behavior would be to turn the phone 90° so that the top of the phone will be held in the user’s left hand and the bottom would be held in the user’s right hand. · The status bar (where the clock, battery power, etc., are found) is hidden when the Game-derived class sets GraphicsDeviceManager.IsFullScreen = true. It is shown when IsFullScreen = false. The default value is false (i.e. the status bar is shown). · You should have a good reason for hiding the status bar. Users find it helpful to know what time it is, how much charge their battery has left, and whether or not their phone is in service range. This is especially true for casual games that you expect someone to play for a few minutes at a time, e.g. while waiting for some event to start, for a phone call to come in, or for a train, bus, or subway to arrive. · In portrait mode, the status bar occupies 32 pixels of space. This means that a game with a back buffer of 480x800 will be scaled down to occupy approximately 461x768 screen pixels. Setting the back buffer to 480x768 (or some resolution with the same 0.625 aspect ratio) will avoid this scaling. · In landscape mode, the status bar occupies 72 pixels of space. This means that a game with a back buffer of 800x480 will be scaled down to occupy approximately 728x437 screen pixels. Setting the back buffer to 728x480 (or some resolution with the same 1.51666667 aspect ratio) will avoid this scaling. Input · Touch input is scaled with screen size. · So if your back buffer is 600x360, a tap in the bottom right corner will come in as (599,359). You don’t need to do anything special to get this automatic scaling of touch behavior. · If you do not use full area of the screen, any touch input outside the area you use will still register as a touch input. For example, if you set a portrait resolution of 240x240, it would be scaled up to occupy a 480x480 area, centered in the screen. If you touch anywhere above this area, you will get a touch input of (X,0) where X is a number from 0 to 239 (in accordance with your 240 pixel wide back buffer). Any touch below this area will give a touch input of (X,239). · If you keep the status bar visible, touches within its area will not be passed to your game. · In general, a screen measurement is the diagonal. So a 3.5” screen is 3.5” long from the bottom right corner to the top left corner. With an aspect ratio of 0.6 (480/800 = 0.6), this means that a phone with a 3.5” screen is only approximately 1.8” wide by 3” tall. So there are approximately 267 pixels in an inch on a 3.5” screen. · Again, this time in metric! 3.5 inches is approximately 8.89 cm. So an 8.89 cm screen is 8.89 cm long from the bottom right corner to the top left corner. With an aspect ratio of 0.6, this means that a phone with an 8.89 cm screen is only approximately 4.57 cm wide by 7.62 cm tall. So there are approximately 105 pixels in a centimeter on an 8.89 cm screen. · Think about the size of your finger tip. If you do not have large hands, think about the size of the fingertip of someone with large hands. Consider that when you are sizing your touch input. Especially consider that when you are spacing two touch targets near one another. You need to judge it for yourself, but items that are next to each other and are each 100x100 should be fine when it comes to selecting items individually. Smaller targets than that are ok provided that you leave space between them. · You want your users to have a pleasant experience. Making touch controls too small or too close to one another will make them nervous about whether they will touch the right target. Take this into account when you plan out your game initially. If possible, do some quick size mockups on an actual phone using colored rectangles that you position and size where you plan to have your game controls. Adjust as necessary. · People do not have transparent hands! Nor are their hands the size of a mouse pointer icon. Consider leaving a dedicated space for input rather than forcing the user to cover up to one-third of the screen with a finger just to play the game. · Another benefit of designing your controls to use a dedicated area is that you’re less likely to have players moving their finger(s) so frantically that they accidentally hit the back button, start button, or search button (many phones have one or more of these on the screen itself – it’s easy to hit one by accident and really annoying if you hit, e.g., the search button and then quickly tap back only to find out that the game didn’t save your progress such that you just wasted all the time you spent playing). · People do not like doing somersaults in order to move something forward with accelerometer-based controls. Test your accelerometer-based controls extensively and get a lot of feedback. Very well-known games from noted publishers have created really bad accelerometer controls and been virtually unplayable as a result. Also be wary of exceptions and other possible failures that the documentation warns about. · When done properly, the accelerometer can add a nice touch to your game (see, e.g. ilomilo where the accelerometer was used to move the background; it added a nice touch without frustrating the user; I also think CarniVale does direct accelerometer controls very well). However, if done poorly, it will make your game an abomination unto the Marketplace. Days, weeks, perhaps even months of development time that you will never get back. I won’t name names; you can search the marketplace for games with terrible reviews and you’ll find them. Graphics · The maximum frame rate is 30 frames per second. This was set as a compromise between battery life and quality. · At least one model of phone is known to have a screen refresh rate that is between 59 and 60 hertz. Because of this, using a fixed time step with a target frame rate of 30 will cause a slight internal delay to build up as the framework is forced to wait slightly for the next refresh. Eventually the delay will get to the point where a draw is skipped in order to recover from the delay. (See Nick's comment below for clarification.) · To deal with that delay, you can either stay with a fixed time step and set the frame rate slightly lower or else you can go to a variable time step and make sure to adjust all of your update data (e.g. player movement distance) to take into account the elapsed time from the last update. A variable time step makes your update logic slightly more complicated but will avoid frame skips entirely. · Currently there are no custom shaders. This might change in the future (there is no hardware limitation preventing it; it simply wasn’t a feature that could be implemented in the time available before launch). · There are five built-in shaders. You can create a lot of nice effects with the built-in shaders. · There is more power on the CPU than there is on the GPU so things you might typically off-load to the GPU will instead make sense to do on the CPU side. · This is a phone. It is not a PC. It is not an Xbox 360. The emulator runs on a PC and uses the full power of your PC. It is very good for testing your code for bugs and doing early prototyping and layout. You should not use it to measure performance. Use actual phone hardware instead. · There are many phone models, each of which has slightly different performance levels for I/O, screen blitting, CPU performance, etc. Do not take your game right to the performance limit on your phone since for some other phones you might be crossing their limits and leaving players with a bad experience. Leave a cushion to account for hardware differences. · Smaller screened phones will have slightly more dots per inch (dpi). Larger screened phones will have slightly less. Either way, the dpi will be much higher than the typical 96 found on most computer screens. Make sure that whoever is doing art for your game takes this into account. · Screens are only required to have 16 bit color (65,536 colors). This is common among smart phones. Using gradients on a 16 bit display can produce an ugly artifact known as banding. Banding is when, rather than a smooth transition from one color to another, you instead see distinct lines. Be careful to avoid this when possible. Banding can be avoided through careful art creation. Its effects can be minimized and even unnoticeable when the texture in question is always moving. You should be careful not to rely on “looks good on my phone” since some phones do have 32-bit displays and thus you’ll find yourself wondering why you’re getting bad reviews that complain about the graphics. Avoid gradients; if you can’t, make sure they are 16-bit safe. Audio · Never rely on sounds as your sole signal to the player that something is happening in the game. They might have the sound off. They might be playing somewhere loud. Etc. · You have to provide controls to disable sound & music. These should be separate. · On at least one model of phone, the volume control API currently has no effect. Players can adjust sound with their hardware volume buttons, but in game selectors simply won’t work. As such, it may not be worth the effort of providing anything beyond on/off switches for sound and music. · MediaPlayer.GameHasControl will return true when a game is hooked up to a PC running Zune. When Zune is running, any attempts to do anything (beyond check GameHasControl) with MediaPlayer will cause an exception to be thrown. If this exception is thrown, catch it and disable music. Exceptions take time to propagate; you don’t want one popping up in every single run of your game’s Update method. · Remember that players can already be listening to music or using the FM radio. In this case GameHasControl will be false and you should handle this appropriately. You can, alternately, ask the player for permission to stop their current music and play your music instead, but the (current) requirement that you restore their music when done is very hard (if not impossible) to deal with. · You can still play sound effects even when the game doesn’t have control of the music, but don’t think this is a backdoor to playing music. Your game will fail certification if your “sound effect” seems to be more like music in scope and length.

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• Installing Oracle 11gR2 on RHEL 6.2

  - by Chris

  Hello all I'm having some difficulty installing Oracle 11gR2 on RHEL 6.2 I have compiled a giant list of every single step I have taken so far I installed RHEL 6.2 on VMWARE it did it's easy install automatically I Selected 4gb of memory Selected max size of 80Gb Selected 2 processors Sorry for the bad styling copy paste isn't working correctly The version of oracle i downloaded is Linux x86-64 11.2.0.1 I am installing this on a local machine NOT a remote machine I followed the following documentation http://docs.oracle.com/cd/E11882_01/install.112/e24326/toc.htm I bolded the steps which I was least sure about from my research Easy installed with RHEL 6.2 for VMWARE Registered with red hat so I can get updates Reinstalled vmware-tools by pressing enter at every choice Sudo yum update at the end something about GPG key selected y then y Checked Memory Requirements grep MemTotal /proc/meminfo MemTotal: 3921368 kb uname -m x86_64 grep SwapTotal /proc/meminfo SwapTotal: 6160376 kb free total used free shared buffers cached Mem: 3921368 2032012 1889356 0 76216 1533268 -/+ buffers/cache: 422528 3498840 Swap: 6160376 0 6160376 df -h /dev/shm Filesystem Size Used Avail Use% Mounted on tmpfs 1.9G 276K 1.9G 1% /dev/shm df -h /tmp Filesystem Size Used Avail Use% Mounted on /dev/sda2 73G 2.7G 67G 4% / df -h Filesystem Size Used Avail Use% Mounted on /dev/sda2 73G 2.7G 67G 4% / tmpfs 1.9G 276K 1.9G 1% /dev/shm /dev/sda1 291M 58M 219M 21% /boot All looked fine to me except maybe for swap? Software Requirements cat /proc/version Linux version 2.6.32-220.el6.x86_64 ([email protected]) (gcc version 4.4.5 20110214 (Red Hat 4.4.5-6) (GCC) ) #1 SMP Wed Nov 9 08:03:13 EST 2011 uname -r 2.6.32-220.el6.x86_64 (same as above but whatever) According to the tutorial should be On Red Hat Enterprise Linux 6 2.6.32-71.el6.x86_64 or later These are the versions of software I have installed binutils-2.20.51.0.2-5.28.el6.x86_64 compat-libcap1-1.10-1.x86_64 compat-libstdc++-33-3.2.3-69.el6.x86_64 compat-libstdc++-33.i686 0:3.2.3-69.el6 gcc-4.4.6-3.el6.x86_64 gcc-c++.x86_64 0:4.4.6-3.el6 glibc-2.12-1.47.el6_2.12.x86_64 glibc-2.12-1.47.el6_2.12.i686 glibc-devel-2.12-1.47.el6_2.12.x86_64 glibc-devel.i686 0:2.12-1.47.el6_2.12 ksh.x86_64 0:20100621-12.el6_2.1 libgcc-4.4.6-3.el6.x86_64 libgcc-4.4.6-3.el6.i686 libstdc++-4.4.6-3.el6.x86_64 libstdc++.i686 0:4.4.6-3.el6 libstdc++-devel.i686 0:4.4.6-3.el6 libstdc++-devel-4.4.6-3.el6.x86_64 libaio-0.3.107-10.el6.x86_64 libaio-0.3.107-10.el6.i686 libaio-devel-0.3.107-10.el6.x86_64 libaio-devel-0.3.107-10.el6.i686 make-3.81-19.el6.x86_64 sysstat-9.0.4-18.el6.x86_64 unixODBC-2.2.14-11.el6.x86_64 unixODBC-devel-2.2.14-11.el6.x86_64 unixODBC-devel-2.2.14-11.el6.i686 unixODBC-2.2.14-11.el6.i686 8. Probably screwed up here or step 9 /usr/sbin/groupadd oinstall /usr/sbin/groupadd dba(not sure why this isn't in the tutorial) /usr/sbin/useradd -g oinstall -G dba oracle passwd oracle /sbin/sysctl -a | grep sem Xkernel.sem = 250 32000 32 128 /sbin/sysctl -a | grep shm kernel.shmmax = 68719476736 kernel.shmall = 4294967296 kernel.shmmni = 4096 vm.hugetlb_shm_group = 0 /sbin/sysctl -a | grep file-max Xfs.file-max = 384629 /sbin/sysctl -a | grep ip_local_port_range Xnet.ipv4.ip_local_port_range = 32768 61000 /sbin/sysctl -a | grep rmem_default Xnet.core.rmem_default = 124928 /sbin/sysctl -a | grep rmem_max Xnet.core.rmem_max = 131071 /sbin/sysctl -a | grep wmem_max Xnet.core.wmem_max = 131071 /sbin/sysctl -a | grep wmem_default Xnet.core.wmem_default = 124928 Here is my sysctl.conf file I only added the items that were bigger: Kernel sysctl configuration file for Red Hat Linux # For binary values, 0 is disabled, 1 is enabled. See sysctl(8) and sysctl.conf(5) for more details. Controls IP packet forwarding net.ipv4.ip_forward = 0 Controls source route verification net.ipv4.conf.default.rp_filter = 1 Do not accept source routing net.ipv4.conf.default.accept_source_route = 0 Controls the System Request debugging functionality of the kernel kernel.sysrq = 0 Controls whether core dumps will append the PID to the core filename. Useful for debugging multi-threaded applications. kernel.core_uses_pid = 1 Controls the use of TCP syncookies net.ipv4.tcp_syncookies = 1 Disable netfilter on bridges. net.bridge.bridge-nf-call-ip6tables = 0 net.bridge.bridge-nf-call-iptables = 0 net.bridge.bridge-nf-call-arptables = 0 Controls the maximum size of a message, in bytes kernel.msgmnb = 65536 Controls the default maxmimum size of a mesage queue kernel.msgmax = 65536 Controls the maximum shared segment size, in bytes kernel.shmmax = 68719476736 Controls the maximum number of shared memory segments, in pages kernel.shmall = 4294967296 fs.aio-max-nr = 1048576 fs.file-max = 6815744 kernel.sem = 250 32000 100 128 net.ipv4.ip_local_port_range = 9000 65500 net.core.rmem_default = 262144 net.core.rmem_max = 4194304 net.core.wmem_default = 262144 net.core.wmem_max = 1048576 /sbin/sysctl -p net.ipv4.ip_forward = 0 net.ipv4.conf.default.rp_filter = 1 net.ipv4.conf.default.accept_source_route = 0 kernel.sysrq = 0 kernel.core_uses_pid = 1 net.ipv4.tcp_syncookies = 1 error: "net.bridge.bridge-nf-call-ip6tables" is an unknown key error: "net.bridge.bridge-nf-call-iptables" is an unknown key error: "net.bridge.bridge-nf-call-arptables" is an unknown key kernel.msgmnb = 65536 kernel.msgmax = 65536 kernel.shmmax = 68719476736 kernel.shmall = 4294967296 fs.aio-max-nr = 1048576 fs.file-max = 6815744 kernel.sem = 250 32000 100 128 net.ipv4.ip_local_port_range = 9000 65500 net.core.rmem_default = 262144 net.core.rmem_max = 4194304 net.core.wmem_default = 262144 net.core.wmem_max = 1048576 su - oracle ulimit -Sn 1024 ulimit -Hn 1024 ulimit -Su 1024 ulimit -Hu 30482 ulimit -Su 1024 ulimit -Ss 10240 ulimit -Hs unlimited su - nano /etc/security/limits.conf *added to the end of the file * oracle soft nproc 2047 oracle hard nproc 16384 oracle soft nofile 1024 oracle hard nofile 65536 oracle soft stack 10240 exit exit su - mkdir -p /app/ chown -R oracle:oinstall /app/ chmod -R 775 /app/ 9. THIS IS PROBABLY WHERE I MESSED UP I then exited out of the root account so now I'm back in my account chris then I su - oracle echo $SHELL /bin/bash umask 0022 (so it should be set already to what is neccesary) Also from what I have read I do not need to set the DISPLAY variable because I'm installing this on the localhost I then opened the .bash_profile of the oracle and changed it to the following .bash_profile Get the aliases and functions if [ -f ~/.bashrc ]; then . ~/.bashrc fi User specific environment and startup programs PATH=$PATH:$HOME/bin; export PATH ORACLE_BASE=/app/oracle ORACLE_SID=orcl export ORACLE_BASE ORACLE_SID I then shutdown the virtual machine shared my desktop folder from my windows 7 then turned back on the virtual machine logged in as chris opened up a terminal then: su - for some reason the shared folder didn't appear so I reinstalled vmware tools again and restarted then same as before su - cp -R linux_oracle/database /db; chown -R oracle:oinstall /db; chmod -R 775 /db; ll /db drwxrwxr-x. 8 oracle oinstall 4096 Jun 5 06:20 database exit su - oracle cd /db/database ./runInstaller AND FINALLY THE INFAMOUS JAVA:132 ERROR MESSAGE Starting Oracle Universal Installer... Checking Temp space: must be greater than 80 MB. Actual 65646 MB Passed Checking swap space: must be greater than 150 MB. Actual 6015 MB Passed Checking monitor: must be configured to display at least 256 colors. Actual 16777216 Passed Preparing to launch Oracle Universal Installer from /tmp/OraInstall2012-06-05_06-47-12AM. Please wait ...[oracle@localhost database]$ Exception in thread "main" java.lang.UnsatisfiedLinkError: /tmp/OraInstall2012-06-05_06-47-12AM/jdk/jre/lib/i386/xawt/libmawt.so: libXext.so.6: cannot open shared object file: No such file or directory at java.lang.ClassLoader$NativeLibrary.load(Native Method) at java.lang.ClassLoader.loadLibrary0(ClassLoader.java:1751) at java.lang.ClassLoader.loadLibrary(ClassLoader.java:1647) at java.lang.Runtime.load0(Runtime.java:769) at java.lang.System.load(System.java:968) at java.lang.ClassLoader$NativeLibrary.load(Native Method) at java.lang.ClassLoader.loadLibrary0(ClassLoader.java:1751) at java.lang.ClassLoader.loadLibrary(ClassLoader.java:1668) at java.lang.Runtime.loadLibrary0(Runtime.java:822) at java.lang.System.loadLibrary(System.java:993) at sun.security.action.LoadLibraryAction.run(LoadLibraryAction.java:50) at java.security.AccessController.doPrivileged(Native Method) at java.awt.Toolkit.loadLibraries(Toolkit.java:1509) at java.awt.Toolkit.(Toolkit.java:1530) at com.jgoodies.looks.LookUtils.isLowResolution(Unknown Source) at com.jgoodies.looks.LookUtils.(Unknown Source) at com.jgoodies.looks.plastic.PlasticLookAndFeel.(PlasticLookAndFeel.java:122) at java.lang.Class.forName0(Native Method) at java.lang.Class.forName(Class.java:242) at javax.swing.SwingUtilities.loadSystemClass(SwingUtilities.java:1783) at javax.swing.UIManager.setLookAndFeel(UIManager.java:480) at oracle.install.commons.util.Application.startup(Application.java:758) at oracle.install.commons.flow.FlowApplication.startup(FlowApplication.java:164) at oracle.install.commons.flow.FlowApplication.startup(FlowApplication.java:181) at oracle.install.commons.base.driver.common.Installer.startup(Installer.java:265) at oracle.install.ivw.db.driver.DBInstaller.startup(DBInstaller.java:114) at oracle.install.ivw.db.driver.DBInstaller.main(DBInstaller.java:132)

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• wcf - maximum array length quota

  - by dav.evans

  Im writing a small wcf/wpf app to resize images but wcf is giving me grief when I try to send an image of size 28K to my service from the client. The service works fine when I send it smaller images. I immediately assumed that this was a configuration issue and I've trawled the web looking at posts regarding the MaxArrayLength property in my binding configuration. Ive upped the limits on these settings on both the client and server to the maximum 2147483647 but still I get the following error: {"The formatter threw an exception while trying to deserialize the message: There was an error while trying to deserialize parameter http://mywebsite.com/services/servicecontracts/2009/01:OriginalImage. The InnerException message was 'There was an error deserializing the object of type System.Drawing.Image. The maximum array length quota (16384) has been exceeded while reading XML data. This quota may be increased by changing the MaxArrayLength property on the XmlDictionaryReaderQuotas object used when creating the XML reader.'. Please see InnerException for more details."} Ive made my client and server configs the same and they look like the following: Server: <system.serviceModel> <bindings> <netTcpBinding> <binding name="NetTcpBinding_ImageResizerServiceContract" closeTimeout="00:01:00" openTimeout="00:01:00" receiveTimeout="00:10:00" sendTimeout="00:01:00" transactionFlow="false" transferMode="Buffered" transactionProtocol="OleTransactions" hostNameComparisonMode="StrongWildcard" listenBacklog="10" maxBufferPoolSize="2147483647" maxBufferSize="2147483647" maxConnections="10" maxReceivedMessageSize="2147483647"> <readerQuotas maxDepth="32" maxStringContentLength="2147483647" maxArrayLength="2147483647" maxBytesPerRead="2147483647" maxNameTableCharCount="2147483647" /> <reliableSession ordered="true" inactivityTimeout="00:10:00" enabled="false" /> <security mode="Transport"> <transport clientCredentialType="Windows" protectionLevel="EncryptAndSign" /> <message clientCredentialType="Windows" /> </security> </binding> </netTcpBinding> </bindings> <behaviors> <serviceBehaviors> <behavior name="ServiceBehavior"> <serviceMetadata httpGetEnabled="true" /> <serviceDebug includeExceptionDetailInFaults="false" /> </behavior> </serviceBehaviors> </behaviors> <services> <service name="LogoResizer.WCF.ServiceTypes.ImageResizerService" behaviorConfiguration="ServiceBehavior"> <host> <baseAddresses> <add baseAddress="http://localhost:900/mex/"/> <add baseAddress="net.tcp://localhost:9000/" /> </baseAddresses> </host> <endpoint binding="netTcpBinding" contract="LogoResizer.WCF.ServiceContracts.IImageResizerService" /> <endpoint address="mex" binding="mexHttpBinding" contract="IMetadataExchange"/> </service> </services> </system.serviceModel> and my client config looks like: <system.serviceModel> <bindings> <netTcpBinding> <binding name="NetTcpBinding_ImageResizerServiceContract" closeTimeout="00:01:00" openTimeout="00:01:00" receiveTimeout="00:10:00" sendTimeout="00:01:00" transactionFlow="false" transferMode="Buffered" transactionProtocol="OleTransactions" hostNameComparisonMode="StrongWildcard" listenBacklog="10" maxBufferPoolSize="2147483647" maxBufferSize="2147483647" maxConnections="10" maxReceivedMessageSize="2147483647"> <readerQuotas maxDepth="32" maxStringContentLength="2147483647" maxArrayLength="2147483647" maxBytesPerRead="2147483647" maxNameTableCharCount="2147483647" /> <reliableSession ordered="true" inactivityTimeout="00:10:00" enabled="false" /> <security mode="Transport"> <transport clientCredentialType="Windows" protectionLevel="EncryptAndSign" /> <message clientCredentialType="Windows" /> </security> </binding> </netTcpBinding> </bindings> <client> <endpoint address="net.tcp://localhost:9000/" binding="netTcpBinding" bindingConfiguration="NetTcpBinding_ImageResizerServiceContract" contract="ImageResizerService.ImageResizerServiceContract" name="NetTcpBinding_ImageResizerServiceContract"> <identity> <userPrincipalName value="[email protected]" /> </identity> </endpoint> </client> </system.serviceModel> It seems no matter what I set these values to I still get an error saying wcf cannot serialize my file because its greater than 16384. Any ideas? edit: the email address in the userPrincipalName tag has been altered for my privacy

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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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• Parsing "true" and "false" using Boost.Spirit.Lex and Boost.Spirit.Qi

  - by Andrew Ross

  As the first stage of a larger grammar using Boost.Spirit I'm trying to parse "true" and "false" to produce the corresponding bool values, true and false. I'm using Spirit.Lex to tokenize the input and have a working implementation for integer and floating point literals (including those expressed in a relaxed scientific notation), exposing int and float attributes. Token definitions #include <boost/spirit/include/lex_lexertl.hpp> namespace lex = boost::spirit::lex; typedef boost::mpl::vector<int, float, bool> token_value_type; template <typename Lexer> struct basic_literal_tokens : lex::lexer<Lexer> { basic_literal_tokens() { this->self.add_pattern("INT", "[-+]?[0-9]+"); int_literal = "{INT}"; // To be lexed as a float a numeric literal must have a decimal point // or include an exponent, otherwise it will be considered an integer. float_literal = "{INT}(((\\.[0-9]+)([eE]{INT})?)|([eE]{INT}))"; literal_true = "true"; literal_false = "false"; this->self = literal_true | literal_false | float_literal | int_literal; } lex::token_def<int> int_literal; lex::token_def<float> float_literal; lex::token_def<bool> literal_true, literal_false; }; Testing parsing of float literals My real implementation uses Boost.Test, but this is a self-contained example. #include <string> #include <iostream> #include <cmath> #include <cstdlib> #include <limits> bool parse_and_check_float(std::string const & input, float expected) { typedef std::string::const_iterator base_iterator_type; typedef lex::lexertl::token<base_iterator_type, token_value_type > token_type; typedef lex::lexertl::lexer<token_type> lexer_type; basic_literal_tokens<lexer_type> basic_literal_lexer; base_iterator_type input_iter(input.begin()); float actual; bool result = lex::tokenize_and_parse(input_iter, input.end(), basic_literal_lexer, basic_literal_lexer.float_literal, actual); return result && std::abs(expected - actual) < std::numeric_limits<float>::epsilon(); } int main(int argc, char *argv[]) { if (parse_and_check_float("+31.4e-1", 3.14)) { return EXIT_SUCCESS; } else { return EXIT_FAILURE; } } Parsing "true" and "false" My problem is when trying to parse "true" and "false". This is the test code I'm using (after removing the Boost.Test parts): bool parse_and_check_bool(std::string const & input, bool expected) { typedef std::string::const_iterator base_iterator_type; typedef lex::lexertl::token<base_iterator_type, token_value_type > token_type; typedef lex::lexertl::lexer<token_type> lexer_type; basic_literal_tokens<lexer_type> basic_literal_lexer; base_iterator_type input_iter(input.begin()); bool actual; lex::token_def<bool> parser = expected ? basic_literal_lexer.literal_true : basic_literal_lexer.literal_false; bool result = lex::tokenize_and_parse(input_iter, input.end(), basic_literal_lexer, parser, actual); return result && actual == expected; } but compilation fails with: boost/spirit/home/qi/detail/assign_to.hpp: In function ‘void boost::spirit::traits::assign_to(const Iterator&, const Iterator&, Attribute&) [with Iterator = __gnu_cxx::__normal_iterator<const char*, std::basic_string<char, std::char_traits<char>, std::allocator<char> > >, Attribute = bool]’: boost/spirit/home/lex/lexer/lexertl/token.hpp:434: instantiated from ‘static void boost::spirit::traits::assign_to_attribute_from_value<Attribute, boost::spirit::lex::lexertl::token<Iterator, AttributeTypes, HasState>, void>::call(const boost::spirit::lex::lexertl::token<Iterator, AttributeTypes, HasState>&, Attribute&) [with Attribute = bool, Iterator = __gnu_cxx::__normal_iterator<const char*, std::basic_string<char, std::char_traits<char>, std::allocator<char> > >, AttributeTypes = boost::mpl::vector<int, float, bool, mpl_::na, mpl_::na, mpl_::na, mpl_::na, mpl_::na, mpl_::na, mpl_::na, mpl_::na, mpl_::na, mpl_::na, mpl_::na, mpl_::na, mpl_::na, mpl_::na, mpl_::na, mpl_::na, mpl_::na>, HasState = mpl_::bool_<true>]’ ... backtrace of instantiation points .... boost/spirit/home/qi/detail/assign_to.hpp:79: error: no matching function for call to ‘boost::spirit::traits::assign_to_attribute_from_iterators<bool, __gnu_cxx::__normal_iterator<const char*, std::basic_string<char, std::char_traits<char>, std::allocator<char> > >, void>::call(const __gnu_cxx::__normal_iterator<const char*, std::basic_string<char, std::char_traits<char>, std::allocator<char> > >&, const __gnu_cxx::__normal_iterator<const char*, std::basic_string<char, std::char_traits<char>, std::allocator<char> > >&, bool&)’ boost/spirit/home/qi/detail/construct.hpp:64: note: candidates are: static void boost::spirit::traits::assign_to_attribute_from_iterators<bool, Iterator, void>::call(const Iterator&, const Iterator&, char&) [with Iterator = __gnu_cxx::__normal_iterator<const char*, std::basic_string<char, std::char_traits<char>, std::allocator<char> > >] My interpretation of this is that Spirit.Qi doesn't know how to convert a string to a bool - surely that's not the case? Has anyone else done this before? If so, how?

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• understand SimpleTimeZone and DST Test

  - by Cygnusx1

  I Have an issue with the use of SimpleTimeZone class in Java. First, the JavaDoc is nice but not quite easy to understand in regards of the start and end Rules. But with the help of some example found on the web, i managed to get it right (i still don't understand why 8 represents the second week of a month in day_of_month!!! but whatever) Now i have written a simple Junit test to validate what i understand: package test; import static org.junit.Assert.assertEquals; import java.sql.Timestamp; import java.util.Calendar; import java.util.GregorianCalendar; import java.util.SimpleTimeZone; import org.apache.log4j.Logger; import org.junit.Test; public class SimpleTimeZoneTest { Logger log = Logger.getLogger(SimpleTimeZoneTest.class); @Test public void testTimeZoneWithDST() throws Exception { Calendar testDateEndOut = new GregorianCalendar(2012, Calendar.NOVEMBER, 4, 01, 59, 59); Calendar testDateEndIn = new GregorianCalendar(2012, Calendar.NOVEMBER, 4, 02, 00, 00); Calendar testDateStartOut = new GregorianCalendar(2012, Calendar.MARCH, 11, 01, 59, 59); Calendar testDateStartIn = new GregorianCalendar(2012, Calendar.MARCH, 11, 02, 00, 00); SimpleTimeZone est = new SimpleTimeZone(-5 * 60 * 60 * 1000, "EST"); est.setStartRule(Calendar.MARCH, 8, -Calendar.SUNDAY, 2 * 60 * 60 * 1000); est.setEndRule(Calendar.NOVEMBER, 1, Calendar.SUNDAY, 2 * 60 * 60 * 1000); Calendar theCal = new GregorianCalendar(est); theCal.setTimeInMillis(testDateEndOut.getTimeInMillis()); log.info(" Cal date = " + new Timestamp(theCal.getTimeInMillis()) + " : " + theCal.getTimeZone().getDisplayName()); log.info(" Cal use DST = " + theCal.getTimeZone().useDaylightTime()); log.info(" Cal In DST = " + theCal.getTimeZone().inDaylightTime(theCal.getTime())); log.info("offset = " + theCal.getTimeZone().getOffset(theCal.getTimeInMillis())); log.info("DTS offset= " + theCal.getTimeZone().getDSTSavings()); assertEquals("End date Should be In DST", true, theCal.getTimeZone().inDaylightTime(theCal.getTime())); theCal.setTimeInMillis(testDateEndIn.getTimeInMillis()); log.info(" Cal date = " + new Timestamp(theCal.getTimeInMillis()) + " : " + theCal.getTimeZone().getDisplayName()); log.info(" Cal use DST = " + theCal.getTimeZone().useDaylightTime()); log.info(" Cal In DST = " + theCal.getTimeZone().inDaylightTime(theCal.getTime())); log.info("offset = " + theCal.getTimeZone().getOffset(theCal.getTimeInMillis())); log.info("DTS offset= " + theCal.getTimeZone().getDSTSavings()); assertEquals("End date Should be Out DST", false, theCal.getTimeZone().inDaylightTime(theCal.getTime())); theCal.setTimeInMillis(testDateStartIn.getTimeInMillis()); log.info(" Cal date = " + new Timestamp(theCal.getTimeInMillis()) + " : " + theCal.getTimeZone().getDisplayName()); log.info(" Cal use DST = " + theCal.getTimeZone().useDaylightTime()); log.info(" Cal In DST = " + theCal.getTimeZone().inDaylightTime(theCal.getTime())); log.info("offset = " + theCal.getTimeZone().getOffset(theCal.getTimeInMillis())); log.info("DTS offset= " + theCal.getTimeZone().getDSTSavings()); assertEquals("Start date Should be in DST", true, theCal.getTimeZone().inDaylightTime(theCal.getTime())); theCal.setTimeInMillis(testDateStartOut.getTimeInMillis()); log.info(" Cal date = " + new Timestamp(theCal.getTimeInMillis()) + " : " + theCal.getTimeZone().getDisplayName()); log.info(" Cal use DST = " + theCal.getTimeZone().useDaylightTime()); log.info(" Cal In DST = " + theCal.getTimeZone().inDaylightTime(theCal.getTime())); log.info("offset = " + theCal.getTimeZone().getOffset(theCal.getTimeInMillis())); log.info("DTS offset= " + theCal.getTimeZone().getDSTSavings()); assertEquals("Start date Should be Out DST", false, theCal.getTimeZone().inDaylightTime(theCal.getTime())); } } Ok, i want to test the date limits to see if the inDaylightTime return the right thing! So, my rules are : DST start the second sunday of March at 2am DST end the first sunday of november at 2am In 2012 (now) this give us the march 11 at 2am and November 4 at 2am You can see my test dates are set properly!!! Well here is the output of my test run: 2012-11-01 18:22:44,344 INFO [test.SimpleTimeZoneTest] - < Cal date = 2012-11-04 01:59:59.0 : Eastern Standard Time> 2012-11-01 18:22:44,345 INFO [test.SimpleTimeZoneTest] - < Cal use DST = true> 2012-11-01 18:22:44,345 INFO [test.SimpleTimeZoneTest] - < Cal In DST = false> 2012-11-01 18:22:44,345 INFO [test.SimpleTimeZoneTest] - <offset = -18000000> 2012-11-01 18:22:44,345 INFO [test.SimpleTimeZoneTest] - <DTS offset= 3600000> My first assert just fails and tell me that 2012-11-04 01:59:59 is not inDST... !!!!??? If i put 2012-11-04 00:59:59, the test pass! This 1 hour gap just puzzle me... can anyone explain this behavior? Oh, btw, if anyone could elaborate on the : est.setStartRule(Calendar.MARCH, 8, -Calendar.SUNDAY, 2 * 60 * 60 * 1000); Why 8 means second week of march... and the -SUNDAY. I can't figure out this thing on a real calendar example!!! Thanks

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• Implementation question involving implementing an interface

  - by Vivin Paliath

  I'm writing a set of collection classes for different types of Trees. I'm doing this as a learning exercise and I'm also hoping it turns out to be something useful. I really want to do this the right way and so I've been reading Effective Java and I've also been looking at the way Joshua Bloch implemented the collection classes by looking at the source. I seem to have a fair idea of what is being done, but I still have a few things to sort out. I have a Node<T> interface and an AbstractNode<T> class that implements the Node interface. I then created a GenericNode<T> (a node that can have 0 to n children, and that is part of an n-ary tree) class that extends AbstractNode<T> and implements Node<T>. This part was easy. Next, I created a Tree<T> interface and an AbstractTree<T> class that implements the Tree<T> interface. After that, I started writing a GenericTree<T> class that extends AbstractTree<T> and implements Tree<T>. This is where I started having problems. As far as the design is concerned, a GenericTree<T> can only consist of nodes of type GenericTreeNode<T>. This includes the root. In my Tree<T> interface I have: public interface Tree<T> { void setRoot(Node<T> root); Node<T> getRoot(); List<Node<T>> postOrder(); ... rest omitted ... } And, AbstractTree<T> implements this interface: public abstract class AbstractTree<T> implements Tree<T> { protected Node<T> root; protected AbstractTree() { } protected AbstractTree(Node<T> root) { this.root = root; } public void setRoot(Node<T> root) { this.root = root; } public Node<T> getRoot() { return this.root; } ... rest omitted ... } In GenericTree<T>, I can have: public GenericTree(Node<T> root) { super(root); } But what this means is that you can create a generic tree using any subtype of Node<T>. You can also set the root of a tree to any subtype of Node<T>. I want to be able to restrict the type of the node to the type of the tree that it can represent. To fix this, I can do this: public GenericTree(GenericNode<T> root) { super(root); } However, setRoot still accepts a parameter of type Node<T>. Which means a user can still create a tree with the wrong type of root node. How do I enforce this constraint? The only way I can think of doing is either: Do an instanceof which limits the check to runtime. I'm not a huge fan of this. Remove setRoot from the interface and have the base class implement this method. This means that it is not part of the contract and anyone who wants to make a new type of tree needs to remember to implement this method. Is there a better way? The second question I have concerns the return type of postOrder which is List<Node<T>>. This means that if a user is operating on a GenericTree<T> object and calls postOrder, he or she receives a list that consists of Node<T> objects. This means when iterating through (using a foreach construct) they would have perform an explicit cast to GenericNode<T> if they want to use methods that are only defined in that class. I don't like having to place this burden on the user. What are my options in this case? I can only think of removing the method from the interface and have the subclass implement this method making sure that it returns a list of appropriate subtype of Node<T>. However, this once again removes it from the contract and it's anyone who wants to create a new type of tree has to remember to implement this method. Is there a better way?

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• WCF 4: Fileless Activation Fails On XP (IIS 5) that has SSL port enabled.

  - by Richard Collette

  I have a service being hosted in IIS on XP via fileless activation. The service starts fine when there is no SSL port enabled for IIS but when the SSL port is enabled, I get the error message: System.ServiceModel.ServiceActivationException: The service '/SkillsPrototype.Web/services/Linkage.svc' cannot be activated due to an exception during compilation. The exception message is: A binding instance has already been associated to listen URI 'http://rcollet.hsb-corp.hsb.com/SkillsPrototype.Web/Services/Linkage.svc'. If two endpoints want to share the same ListenUri, they must also share the same binding object instance. The two conflicting endpoints were either specified in AddServiceEndpoint() calls, in a config file, or a combination of AddServiceEndpoint() and config. . ---> System.InvalidOperationException: A binding instance has already been associated to listen URI 'http://rcollet.hsb-corp.hsb.com/SkillsPrototype.Web/Services/Linkage.svc'. If two endpoints want to share the same ListenUri, they must also share the same binding object instance. The two conflicting endpoints were either specified in AddServiceEndpoint() calls, in a config file, or a combination of AddServiceEndpoint() and config. My service model configuration is <system.serviceModel> <diagnostics wmiProviderEnabled="true"> <messageLogging logEntireMessage="true" logMalformedMessages="true" logMessagesAtServiceLevel="true" logMessagesAtTransportLevel="true" maxMessagesToLog="3000"/> </diagnostics> <standardEndpoints> <webHttpEndpoint> <standardEndpoint name="" helpEnabled="true" automaticFormatSelectionEnabled="true" /> </webHttpEndpoint> </standardEndpoints> <behaviors> <serviceBehaviors> <behavior> <serviceMetadata httpGetEnabled="true"/> <serviceDebug includeExceptionDetailInFaults="true" /> </behavior> </serviceBehaviors> </behaviors> <bindings> <webHttpBinding> <binding> <security mode="None"> <transport clientCredentialType="None"/> </security> </binding> </webHttpBinding> </bindings> <protocolMapping> </protocolMapping> <services> </services> <serviceHostingEnvironment multipleSiteBindingsEnabled="false"> <serviceActivations> <clear/> <add factory="System.ServiceModel.Activation.WebScriptServiceHostFactory" service="SkillsPrototype.ServiceModel.Linkage" relativeAddress="~/Services/Linkage.svc"/> </serviceActivations> </serviceHostingEnvironment> </system.serviceModel> When you look in the svclog file, there two base addresses that are returned when SSL is enabled, one for http and one for https. I suspect that this is part of the issue but I am not sure how to resolve it. <E2ETraceEvent xmlns="http://schemas.microsoft.com/2004/06/E2ETraceEvent"> <System xmlns="http://schemas.microsoft.com/2004/06/windows/eventlog/system"> <EventID>524333</EventID> <Type>3</Type> <SubType Name="Information">0</SubType> <Level>8</Level> <TimeCreated SystemTime="2010-06-16T17:40:55.8168605Z" /> <Source Name="System.ServiceModel" /> <Correlation ActivityID="{95927f9a-fa90-46f4-af8b-721322a87aaa}" /> <Execution ProcessName="aspnet_wp" ProcessID="1888" ThreadID="5" /> <Channel/> <Computer>RCOLLET</Computer> </System> <ApplicationData> <TraceData> <DataItem> <TraceRecord xmlns="http://schemas.microsoft.com/2004/10/E2ETraceEvent/TraceRecord" Severity="Information"> <TraceIdentifier>http://msdn.microsoft.com/en-US/library/System.ServiceModel.ServiceHostBaseAddresses.aspx</TraceIdentifier> <Description>ServiceHost base addresses.</Description> <AppDomain>/LM/w3svc/1/ROOT/SkillsPrototype.Web-1-129211836532542949</AppDomain> <Source>System.ServiceModel.WebScriptServiceHost/49153359</Source> <ExtendedData xmlns="http://schemas.microsoft.com/2006/08/ServiceModel/CollectionTraceRecord"> <BaseAddresses> <Address>http://rcollet.hsb-corp.hsb.com/SkillsPrototype.Web/Services/Linkage.svc</Address> <Address>https://rcollet.hsb-corp.hsb.com/SkillsPrototype.Web/Services/Linkage.svc</Address> </BaseAddresses> </ExtendedData> </TraceRecord> </DataItem> </TraceData> </ApplicationData> </E2ETraceEvent> I can't post the full service log due to character limits on the post.

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• JSF : How to refresh required field in ajax request

  - by Tama

  Ok, here you are the core problem. The page. I have two required "input text". A command button that changes the bean value and reRenderes the "job" object. <a4j:form id="pervForm"> SURNAME:<h:inputText id="surname" label="Surname" value="#{prevManager.surname}" required="true" /> <br/> JOB:<h:inputText value="#{prevManager.job}" id="job" maxlength="10" size="10" label="#{msg.common_label_job}" required="true" /> <br/> <a4j:commandButton value="Set job to Programmer" ajaxSingle="true" reRender="job"> <a4j:actionparam name="jVal" value="Programmer" assignTo="#{prevManager.job}"/> </a4j:commandButton> <h:commandButton id="save" value="save" action="save" class="HATSBUTTON"/> </a4j:form> Here the simple manager: public class PrevManager { private String surname; private String job; public String getSurname() { return surname; } public void setSurname(String surname) { this.surname = surname; } public String getJob() { return job; } public void setJob(String job) { this.job = job; } public String save() { //do something } } Let's do this: Write something on the Job input text (such as "teacher"). Leave empty the surname. Save. Validation error appears (surname is mandatory). Press "Set job to Programmer": nothing happens. Checking the bean value, I discovered that it is correctly updated, indeed the component on the page is not updated! Well, according to the JBoss Docs I found: Ajax region is a key ajax component. It limits the part of the component tree to be processed on the server side when ajax request comes. Processing means invocation during Decode, Validation and Model Update phase. Most common reasons to use a region are: -avoiding the aborting of the JSF lifecycle processing during the validation of other form input unnecessary for given ajax request; -defining the different strategies when events will be delivered (immediate="true/false") -showing an individual indicator of an ajax status -increasing the performance of the rendering processing (selfRendered="true/false", renderRegionOnly="true/false") The following two examples show the situation when a validation error does not allow to process an ajax input. Type the name. The outputText component should reappear after you. However, in the first case, this activity will be aborted because of the other field with required="true". You will see only the error message while the "Job" field is empty. Here you are the example: <ui:composition xmlns="http://www.w3.org/1999/xhtml" xmlns:ui="http://java.sun.com/jsf/facelets" xmlns:h="http://java.sun.com/jsf/html" xmlns:f="http://java.sun.com/jsf/core" xmlns:a4j="http://richfaces.org/a4j" xmlns:rich="http://richfaces.org/rich"> <style> .outergridvalidationcolumn { padding: 0px 30px 10px 0px; } </style> <a4j:outputPanel ajaxRendered="true"> <h:messages style="color:red" /> </a4j:outputPanel> <h:panelGrid columns="2" columnClasses="outergridvalidationcolumn"> <h:form id="form1"> <h:panelGrid columns="2"> <h:outputText value="Name" /> <h:inputText value="#{userBean.name}"> <a4j:support event="onkeyup" reRender="outname" /> </h:inputText> <h:outputText value="Job" /> <h:inputText required="true" id="job2" value="#{userBean.job}" /> </h:panelGrid> </h:form> <h:form id="form2"> <h:panelGrid columns="2"> <h:outputText value="Name" /> <a4j:region> <h:inputText value="#{userBean.name}"> <a4j:support event="onkeyup" reRender="outname" /> </h:inputText> </a4j:region> <h:outputText value="Job" /> <h:inputText required="true" id="job1" value="#{userBean.job}" /> </h:panelGrid> </h:form> </h:panelGrid> <h:outputText id="outname" style="font-weight:bold" value="Typed Name: #{userBean.name}" /> <br /> </ui:composition> Form1: the behaviour is incorrect. I need to fill the job and then the name. Form2: the behaviour is correct. I do not need to fill the job to see the correct value. Unfortunately using Ajax region does not help (indeed I used it in a bad way ...) because my fields are both REQUIRED. That's the main different. Any idea? Many thanks.

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• Trying to reduce the speed overhead of an almost-but-not-quite-int number class

  - by Fumiyo Eda

  I have implemented a C++ class which behaves very similarly to the standard int type. The difference is that it has an additional concept of "epsilon" which represents some tiny value that is much less than 1, but greater than 0. One way to think of it is as a very wide fixed point number with 32 MSBs (the integer parts), 32 LSBs (the epsilon parts) and a huge sea of zeros in between. The following class works, but introduces a ~2x speed penalty in the overall program. (The program includes code that has nothing to do with this class, so the actual speed penalty of this class is probably much greater than 2x.) I can't paste the code that is using this class, but I can say the following: +, -, +=, <, > and >= are the only heavily used operators. Use of setEpsilon() and getInt() is extremely rare. * is also rare, and does not even need to consider the epsilon values at all. Here is the class: #include <limits> struct int32Uepsilon { typedef int32Uepsilon Self; int32Uepsilon () { _value = 0; _eps = 0; } int32Uepsilon (const int &i) { _value = i; _eps = 0; } void setEpsilon() { _eps = 1; } Self operator+(const Self &rhs) const { Self result = *this; result._value += rhs._value; result._eps += rhs._eps; return result; } Self operator-(const Self &rhs) const { Self result = *this; result._value -= rhs._value; result._eps -= rhs._eps; return result; } Self operator-( ) const { Self result = *this; result._value = -result._value; result._eps = -result._eps; return result; } Self operator*(const Self &rhs) const { return this->getInt() * rhs.getInt(); } // XXX: discards epsilon bool operator<(const Self &rhs) const { return (_value < rhs._value) || (_value == rhs._value && _eps < rhs._eps); } bool operator>(const Self &rhs) const { return (_value > rhs._value) || (_value == rhs._value && _eps > rhs._eps); } bool operator>=(const Self &rhs) const { return (_value >= rhs._value) || (_value == rhs._value && _eps >= rhs._eps); } Self &operator+=(const Self &rhs) { this->_value += rhs._value; this->_eps += rhs._eps; return *this; } Self &operator-=(const Self &rhs) { this->_value -= rhs._value; this->_eps -= rhs._eps; return *this; } int getInt() const { return(_value); } private: int _value; int _eps; }; namespace std { template<> struct numeric_limits<int32Uepsilon> { static const bool is_signed = true; static int max() { return 2147483647; } } }; The code above works, but it is quite slow. Does anyone have any ideas on how to improve performance? There are a few hints/details I can give that might be helpful: 32 bits are definitely insufficient to hold both _value and _eps. In practice, up to 24 ~ 28 bits of _value are used and up to 20 bits of _eps are used. I could not measure a significant performance difference between using int32_t and int64_t, so memory overhead itself is probably not the problem here. Saturating addition/subtraction on _eps would be cool, but isn't really necessary. Note that the signs of _value and _eps are not necessarily the same! This broke my first attempt at speeding this class up. Inline assembly is no problem, so long as it works with GCC on a Core i7 system running Linux!

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• Loop crashing program having to do with 2D arrays

  - by user450062

  I am creating an encoding program and when I instruct the program to create a 5X5 grid based on the alphabet while skipping over letters that match up to certain pre-defined variables(which are given values by user input during runtime). I have a loop that instructs the loop to keep running until the values that access the array are out of bounds, the loop seems to cause the problem. This code is standardized so there shouldn't be much trouble compiling it in another compiler. Also would it be better to seperate my program into functions? here is the code: #include<iostream> #include<fstream> #include<cstdlib> #include<string> #include<limits> using namespace std; int main(){ while (!cin.fail()) { char type[81]; char filename[20]; char key [5]; char f[2] = "q"; char g[2] = "q"; char h[2] = "q"; char i[2] = "q"; char j[2] = "q"; char k[2] = "q"; char l[2] = "q"; int a = 1; int b = 1; int c = 1; int d = 1; int e = 1; string cipherarraytemplate[5][5]= { {"a","b","c","d","e"}, {"f","g","h","i","j"}, {"k","l","m","n","o"}, {"p","r","s","t","u"}, {"v","w","x","y","z"} }; string cipherarray[5][5]= { {"a","b","c","d","e"}, {"f","g","h","i","j"}, {"k","l","m","n","o"}, {"p","r","s","t","u"}, {"v","w","x","y","z"} }; cout<<"Enter the name of a file you want to create.\n"; cin>>filename; ofstream outFile; outFile.open(filename); outFile<<fixed; outFile.precision(2); outFile.setf(ios_base::showpoint); cin.ignore(std::numeric_limits<int>::max(),'\n'); cout<<"enter your codeword(codeword can have no repeating letters)\n"; cin>>key; while (key[a] != '\0' ){ while(b < 6){ cipherarray[b][c] = key[a]; if ( f == "q" ) { cipherarray[b][c] = f; } if ( f != "q" && g == "q" ) { cipherarray[b][c] = g; } if ( g != "q" && h == "q" ) { cipherarray[b][c] = h; } if ( h != "q" && i == "q" ) { cipherarray[b][c] = i; } if ( i != "q" && j == "q" ) { cipherarray[b][c] = j; } if ( j != "q" && k == "q" ) { cipherarray[b][c] = k; } if ( k != "q" && l == "q" ) { cipherarray[b][c] = l; } a++; b++; } c++; b = 1; } while (c < 6 || b < 6){ if (cipherarraytemplate[d][e] == f || cipherarraytemplate[d][e] == g || cipherarraytemplate[d][e] == h || cipherarraytemplate[d][e] == i || cipherarraytemplate[d][e] == j || cipherarraytemplate[d][e] == k || cipherarraytemplate[d][e] == l){ d++; } else { cipherarray[b][c] = cipherarraytemplate[d][e]; d++; b++; } if (d == 6){ d = 1; e++; } if (b == 6){ c++; b = 1; } } cout<<"now enter some text."<<endl<<"To end this program press Crtl-Z\n"; while(!cin.fail()){ cin.getline(type,81); outFile<<type<<endl; } outFile.close(); } } I know there is going to be some mid-forties guy out there who is going to stumble on to this post, he's have been programming for 20-some years and he's going to look at my code and say: "what is this guy doing".

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• A Taxonomy of Numerical Methods v1

  - by JoshReuben

  Numerical Analysis – When, What, (but not how) Once you understand the Math & know C++, Numerical Methods are basically blocks of iterative & conditional math code. I found the real trick was seeing the forest for the trees – knowing which method to use for which situation. Its pretty easy to get lost in the details – so I’ve tried to organize these methods in a way that I can quickly look this up. I’ve included links to detailed explanations and to C++ code examples. I’ve tried to classify Numerical methods in the following broad categories: Solving Systems of Linear Equations Solving Non-Linear Equations Iteratively Interpolation Curve Fitting Optimization Numerical Differentiation & Integration Solving ODEs Boundary Problems Solving EigenValue problems Enjoy – I did ! Solving Systems of Linear Equations Overview Solve sets of algebraic equations with x unknowns The set is commonly in matrix form Gauss-Jordan Elimination http://en.wikipedia.org/wiki/Gauss%E2%80%93Jordan_elimination C++: http://www.codekeep.net/snippets/623f1923-e03c-4636-8c92-c9dc7aa0d3c0.aspx Produces solution of the equations & the coefficient matrix Efficient, stable 2 steps: · Forward Elimination – matrix decomposition: reduce set to triangular form (0s below the diagonal) or row echelon form. If degenerate, then there is no solution · Backward Elimination –write the original matrix as the product of ints inverse matrix & its reduced row-echelon matrix à reduce set to row canonical form & use back-substitution to find the solution to the set Elementary ops for matrix decomposition: · Row multiplication · Row switching · Add multiples of rows to other rows Use pivoting to ensure rows are ordered for achieving triangular form LU Decomposition http://en.wikipedia.org/wiki/LU_decomposition C++: http://ganeshtiwaridotcomdotnp.blogspot.co.il/2009/12/c-c-code-lu-decomposition-for-solving.html Represent the matrix as a product of lower & upper triangular matrices A modified version of GJ Elimination Advantage – can easily apply forward & backward elimination to solve triangular matrices Techniques: · Doolittle Method – sets the L matrix diagonal to unity · Crout Method - sets the U matrix diagonal to unity Note: both the L & U matrices share the same unity diagonal & can be stored compactly in the same matrix Gauss-Seidel Iteration http://en.wikipedia.org/wiki/Gauss%E2%80%93Seidel_method C++: http://www.nr.com/forum/showthread.php?t=722 Transform the linear set of equations into a single equation & then use numerical integration (as integration formulas have Sums, it is implemented iteratively). an optimization of Gauss-Jacobi: 1.5 times faster, requires 0.25 iterations to achieve the same tolerance Solving Non-Linear Equations Iteratively find roots of polynomials – there may be 0, 1 or n solutions for an n order polynomial use iterative techniques Iterative methods · used when there are no known analytical techniques · Requires set functions to be continuous & differentiable · Requires an initial seed value – choice is critical to convergence à conduct multiple runs with different starting points & then select best result · Systematic - iterate until diminishing returns, tolerance or max iteration conditions are met · bracketing techniques will always yield convergent solutions, non-bracketing methods may fail to converge Incremental method if a nonlinear function has opposite signs at 2 ends of a small interval x1 & x2, then there is likely to be a solution in their interval – solutions are detected by evaluating a function over interval steps, for a change in sign, adjusting the step size dynamically. Limitations – can miss closely spaced solutions in large intervals, cannot detect degenerate (coinciding) solutions, limited to functions that cross the x-axis, gives false positives for singularities Fixed point method http://en.wikipedia.org/wiki/Fixed-point_iteration C++: http://books.google.co.il/books?id=weYj75E_t6MC&pg=PA79&lpg=PA79&dq=fixed+point+method++c%2B%2B&source=bl&ots=LQ-5P_taoC&sig=lENUUIYBK53tZtTwNfHLy5PEWDk&hl=en&sa=X&ei=wezDUPW1J5DptQaMsIHQCw&redir_esc=y#v=onepage&q=fixed%20point%20method%20%20c%2B%2B&f=false Algebraically rearrange a solution to isolate a variable then apply incremental method Bisection method http://en.wikipedia.org/wiki/Bisection_method C++: http://numericalcomputing.wordpress.com/category/algorithms/ Bracketed - Select an initial interval, keep bisecting it ad midpoint into sub-intervals and then apply incremental method on smaller & smaller intervals – zoom in Adv: unaffected by function gradient à reliable Disadv: slow convergence False Position Method http://en.wikipedia.org/wiki/False_position_method C++: http://www.dreamincode.net/forums/topic/126100-bisection-and-false-position-methods/ Bracketed - Select an initial interval , & use the relative value of function at interval end points to select next sub-intervals (estimate how far between the end points the solution might be & subdivide based on this) Newton-Raphson method http://en.wikipedia.org/wiki/Newton's_method C++: http://www-users.cselabs.umn.edu/classes/Summer-2012/csci1113/index.php?page=./newt3 Also known as Newton's method Convenient, efficient Not bracketed – only a single initial guess is required to start iteration – requires an analytical expression for the first derivative of the function as input. Evaluates the function & its derivative at each step. Can be extended to the Newton MutiRoot method for solving multiple roots Can be easily applied to an of n-coupled set of non-linear equations – conduct a Taylor Series expansion of a function, dropping terms of order n, rewrite as a Jacobian matrix of PDs & convert to simultaneous linear equations !!! Secant Method http://en.wikipedia.org/wiki/Secant_method C++: http://forum.vcoderz.com/showthread.php?p=205230 Unlike N-R, can estimate first derivative from an initial interval (does not require root to be bracketed) instead of inputting it Since derivative is approximated, may converge slower. Is fast in practice as it does not have to evaluate the derivative at each step. Similar implementation to False Positive method Birge-Vieta Method http://mat.iitm.ac.in/home/sryedida/public_html/caimna/transcendental/polynomial%20methods/bv%20method.html C++: http://books.google.co.il/books?id=cL1boM2uyQwC&pg=SA3-PA51&lpg=SA3-PA51&dq=Birge-Vieta+Method+c%2B%2B&source=bl&ots=QZmnDTK3rC&sig=BPNcHHbpR_DKVoZXrLi4nVXD-gg&hl=en&sa=X&ei=R-_DUK2iNIjzsgbE5ID4Dg&redir_esc=y#v=onepage&q=Birge-Vieta%20Method%20c%2B%2B&f=false combines Horner's method of polynomial evaluation (transforming into lesser degree polynomials that are more computationally efficient to process) with Newton-Raphson to provide a computational speed-up Interpolation Overview Construct new data points for as close as possible fit within range of a discrete set of known points (that were obtained via sampling, experimentation) Use Taylor Series Expansion of a function f(x) around a specific value for x Linear Interpolation http://en.wikipedia.org/wiki/Linear_interpolation C++: http://www.hamaluik.com/?p=289 Straight line between 2 points à concatenate interpolants between each pair of data points Bilinear Interpolation http://en.wikipedia.org/wiki/Bilinear_interpolation C++: http://supercomputingblog.com/graphics/coding-bilinear-interpolation/2/ Extension of the linear function for interpolating functions of 2 variables – perform linear interpolation first in 1 direction, then in another. Used in image processing – e.g. texture mapping filter. Uses 4 vertices to interpolate a value within a unit cell. Lagrange Interpolation http://en.wikipedia.org/wiki/Lagrange_polynomial C++: http://www.codecogs.com/code/maths/approximation/interpolation/lagrange.php For polynomials Requires recomputation for all terms for each distinct x value – can only be applied for small number of nodes Numerically unstable Barycentric Interpolation http://epubs.siam.org/doi/pdf/10.1137/S0036144502417715 C++: http://www.gamedev.net/topic/621445-barycentric-coordinates-c-code-check/ Rearrange the terms in the equation of the Legrange interpolation by defining weight functions that are independent of the interpolated value of x Newton Divided Difference Interpolation http://en.wikipedia.org/wiki/Newton_polynomial C++: http://jee-appy.blogspot.co.il/2011/12/newton-divided-difference-interpolation.html Hermite Divided Differences: Interpolation polynomial approximation for a given set of data points in the NR form - divided differences are used to approximately calculate the various differences. For a given set of 3 data points , fit a quadratic interpolant through the data Bracketed functions allow Newton divided differences to be calculated recursively Difference table Cubic Spline Interpolation http://en.wikipedia.org/wiki/Spline_interpolation C++: https://www.marcusbannerman.co.uk/index.php/home/latestarticles/42-articles/96-cubic-spline-class.html Spline is a piecewise polynomial Provides smoothness – for interpolations with significantly varying data Use weighted coefficients to bend the function to be smooth & its 1st & 2nd derivatives are continuous through the edge points in the interval Curve Fitting A generalization of interpolating whereby given data points may contain noise à the curve does not necessarily pass through all the points Least Squares Fit http://en.wikipedia.org/wiki/Least_squares C++: http://www.ccas.ru/mmes/educat/lab04k/02/least-squares.c Residual – difference between observed value & expected value Model function is often chosen as a linear combination of the specified functions Determines: A) The model instance in which the sum of squared residuals has the least value B) param values for which model best fits data Straight Line Fit Linear correlation between independent variable and dependent variable Linear Regression http://en.wikipedia.org/wiki/Linear_regression C++: http://www.oocities.org/david_swaim/cpp/linregc.htm Special case of statistically exact extrapolation Leverage least squares Given a basis function, the sum of the residuals is determined and the corresponding gradient equation is expressed as a set of normal linear equations in matrix form that can be solved (e.g. using LU Decomposition) Can be weighted - Drop the assumption that all errors have the same significance –-> confidence of accuracy is different for each data point. Fit the function closer to points with higher weights Polynomial Fit - use a polynomial basis function Moving Average http://en.wikipedia.org/wiki/Moving_average C++: http://www.codeproject.com/Articles/17860/A-Simple-Moving-Average-Algorithm Used for smoothing (cancel fluctuations to highlight longer-term trends & cycles), time series data analysis, signal processing filters Replace each data point with average of neighbors. Can be simple (SMA), weighted (WMA), exponential (EMA). Lags behind latest data points – extra weight can be given to more recent data points. Weights can decrease arithmetically or exponentially according to distance from point. Parameters: smoothing factor, period, weight basis Optimization Overview Given function with multiple variables, find Min (or max by minimizing –f(x)) Iterative approach Efficient, but not necessarily reliable Conditions: noisy data, constraints, non-linear models Detection via sign of first derivative - Derivative of saddle points will be 0 Local minima Bisection method Similar method for finding a root for a non-linear equation Start with an interval that contains a minimum Golden Search method http://en.wikipedia.org/wiki/Golden_section_search C++: http://www.codecogs.com/code/maths/optimization/golden.php Bisect intervals according to golden ratio 0.618.. Achieves reduction by evaluating a single function instead of 2 Newton-Raphson Method Brent method http://en.wikipedia.org/wiki/Brent's_method C++: http://people.sc.fsu.edu/~jburkardt/cpp_src/brent/brent.cpp Based on quadratic or parabolic interpolation – if the function is smooth & parabolic near to the minimum, then a parabola fitted through any 3 points should approximate the minima – fails when the 3 points are collinear , in which case the denominator is 0 Simplex Method http://en.wikipedia.org/wiki/Simplex_algorithm C++: http://www.codeguru.com/cpp/article.php/c17505/Simplex-Optimization-Algorithm-and-Implemetation-in-C-Programming.htm Find the global minima of any multi-variable function Direct search – no derivatives required At each step it maintains a non-degenerative simplex – a convex hull of n+1 vertices. Obtains the minimum for a function with n variables by evaluating the function at n-1 points, iteratively replacing the point of worst result with the point of best result, shrinking the multidimensional simplex around the best point. Point replacement involves expanding & contracting the simplex near the worst value point to determine a better replacement point Oscillation can be avoided by choosing the 2nd worst result Restart if it gets stuck Parameters: contraction & expansion factors Simulated Annealing http://en.wikipedia.org/wiki/Simulated_annealing C++: http://code.google.com/p/cppsimulatedannealing/ Analogy to heating & cooling metal to strengthen its structure Stochastic method – apply random permutation search for global minima - Avoid entrapment in local minima via hill climbing Heating schedule - Annealing schedule params: temperature, iterations at each temp, temperature delta Cooling schedule – can be linear, step-wise or exponential Differential Evolution http://en.wikipedia.org/wiki/Differential_evolution C++: http://www.amichel.com/de/doc/html/ More advanced stochastic methods analogous to biological processes: Genetic algorithms, evolution strategies Parallel direct search method against multiple discrete or continuous variables Initial population of variable vectors chosen randomly – if weighted difference vector of 2 vectors yields a lower objective function value then it replaces the comparison vector Many params: #parents, #variables, step size, crossover constant etc Convergence is slow – many more function evaluations than simulated annealing Numerical Differentiation Overview 2 approaches to finite difference methods: · A) approximate function via polynomial interpolation then differentiate · B) Taylor series approximation – additionally provides error estimate Finite Difference methods http://en.wikipedia.org/wiki/Finite_difference_method C++: http://www.wpi.edu/Pubs/ETD/Available/etd-051807-164436/unrestricted/EAMPADU.pdf Find differences between high order derivative values - Approximate differential equations by finite differences at evenly spaced data points Based on forward & backward Taylor series expansion of f(x) about x plus or minus multiples of delta h. Forward / backward difference - the sums of the series contains even derivatives and the difference of the series contains odd derivatives – coupled equations that can be solved. Provide an approximation of the derivative within a O(h^2) accuracy There is also central difference & extended central difference which has a O(h^4) accuracy Richardson Extrapolation http://en.wikipedia.org/wiki/Richardson_extrapolation C++: http://mathscoding.blogspot.co.il/2012/02/introduction-richardson-extrapolation.html A sequence acceleration method applied to finite differences Fast convergence, high accuracy O(h^4) Derivatives via Interpolation Cannot apply Finite Difference method to discrete data points at uneven intervals – so need to approximate the derivative of f(x) using the derivative of the interpolant via 3 point Lagrange Interpolation Note: the higher the order of the derivative, the lower the approximation precision Numerical Integration Estimate finite & infinite integrals of functions More accurate procedure than numerical differentiation Use when it is not possible to obtain an integral of a function analytically or when the function is not given, only the data points are Newton Cotes Methods http://en.wikipedia.org/wiki/Newton%E2%80%93Cotes_formulas C++: http://www.siafoo.net/snippet/324 For equally spaced data points Computationally easy – based on local interpolation of n rectangular strip areas that is piecewise fitted to a polynomial to get the sum total area Evaluate the integrand at n+1 evenly spaced points – approximate definite integral by Sum Weights are derived from Lagrange Basis polynomials Leverage Trapezoidal Rule for default 2nd formulas, Simpson 1/3 Rule for substituting 3 point formulas, Simpson 3/8 Rule for 4 point formulas. For 4 point formulas use Bodes Rule. Higher orders obtain more accurate results Trapezoidal Rule uses simple area, Simpsons Rule replaces the integrand f(x) with a quadratic polynomial p(x) that uses the same values as f(x) for its end points, but adds a midpoint Romberg Integration http://en.wikipedia.org/wiki/Romberg's_method C++: http://code.google.com/p/romberg-integration/downloads/detail?name=romberg.cpp&can=2&q= Combines trapezoidal rule with Richardson Extrapolation Evaluates the integrand at equally spaced points The integrand must have continuous derivatives Each R(n,m) extrapolation uses a higher order integrand polynomial replacement rule (zeroth starts with trapezoidal) à a lower triangular matrix set of equation coefficients where the bottom right term has the most accurate approximation. The process continues until the difference between 2 successive diagonal terms becomes sufficiently small. Gaussian Quadrature http://en.wikipedia.org/wiki/Gaussian_quadrature C++: http://www.alglib.net/integration/gaussianquadratures.php Data points are chosen to yield best possible accuracy – requires fewer evaluations Ability to handle singularities, functions that are difficult to evaluate The integrand can include a weighting function determined by a set of orthogonal polynomials. Points & weights are selected so that the integrand yields the exact integral if f(x) is a polynomial of degree <= 2n+1 Techniques (basically different weighting functions): · Gauss-Legendre Integration w(x)=1 · Gauss-Laguerre Integration w(x)=e^-x · Gauss-Hermite Integration w(x)=e^-x^2 · Gauss-Chebyshev Integration w(x)= 1 / Sqrt(1-x^2) Solving ODEs Use when high order differential equations cannot be solved analytically Evaluated under boundary conditions RK for systems – a high order differential equation can always be transformed into a coupled first order system of equations Euler method http://en.wikipedia.org/wiki/Euler_method C++: http://rosettacode.org/wiki/Euler_method First order Runge–Kutta method. Simple recursive method – given an initial value, calculate derivative deltas. Unstable & not very accurate (O(h) error) – not used in practice A first-order method - the local error (truncation error per step) is proportional to the square of the step size, and the global error (error at a given time) is proportional to the step size In evolving solution between data points xn & xn+1, only evaluates derivatives at beginning of interval xn à asymmetric at boundaries Higher order Runge Kutta http://en.wikipedia.org/wiki/Runge%E2%80%93Kutta_methods C++: http://www.dreamincode.net/code/snippet1441.htm 2nd & 4th order RK - Introduces parameterized midpoints for more symmetric solutions à accuracy at higher computational cost Adaptive RK – RK-Fehlberg – estimate the truncation at each integration step & automatically adjust the step size to keep error within prescribed limits. At each step 2 approximations are compared – if in disagreement to a specific accuracy, the step size is reduced Boundary Value Problems Where solution of differential equations are located at 2 different values of the independent variable x à more difficult, because cannot just start at point of initial value – there may not be enough starting conditions available at the end points to produce a unique solution An n-order equation will require n boundary conditions – need to determine the missing n-1 conditions which cause the given conditions at the other boundary to be satisfied Shooting Method http://en.wikipedia.org/wiki/Shooting_method C++: http://ganeshtiwaridotcomdotnp.blogspot.co.il/2009/12/c-c-code-shooting-method-for-solving.html Iteratively guess the missing values for one end & integrate, then inspect the discrepancy with the boundary values of the other end to adjust the estimate Given the starting boundary values u1 & u2 which contain the root u, solve u given the false position method (solving the differential equation as an initial value problem via 4th order RK), then use u to solve the differential equations. Finite Difference Method For linear & non-linear systems Higher order derivatives require more computational steps – some combinations for boundary conditions may not work though Improve the accuracy by increasing the number of mesh points Solving EigenValue Problems An eigenvalue can substitute a matrix when doing matrix multiplication à convert matrix multiplication into a polynomial EigenValue For a given set of equations in matrix form, determine what are the solution eigenvalue & eigenvectors Similar Matrices - have same eigenvalues. Use orthogonal similarity transforms to reduce a matrix to diagonal form from which eigenvalue(s) & eigenvectors can be computed iteratively Jacobi method http://en.wikipedia.org/wiki/Jacobi_method C++: http://people.sc.fsu.edu/~jburkardt/classes/acs2_2008/openmp/jacobi/jacobi.html Robust but Computationally intense – use for small matrices < 10x10 Power Iteration http://en.wikipedia.org/wiki/Power_iteration For any given real symmetric matrix, generate the largest single eigenvalue & its eigenvectors Simplest method – does not compute matrix decomposition à suitable for large, sparse matrices Inverse Iteration Variation of power iteration method – generates the smallest eigenvalue from the inverse matrix Rayleigh Method http://en.wikipedia.org/wiki/Rayleigh's_method_of_dimensional_analysis Variation of power iteration method Rayleigh Quotient Method Variation of inverse iteration method Matrix Tri-diagonalization Method Use householder algorithm to reduce an NxN symmetric matrix to a tridiagonal real symmetric matrix vua N-2 orthogonal transforms     Whats Next Outside of Numerical Methods there are lots of different types of algorithms that I’ve learned over the decades: Data Mining – (I covered this briefly in a previous post: http://geekswithblogs.net/JoshReuben/archive/2007/12/31/ssas-dm-algorithms.aspx ) Search & Sort Routing Problem Solving Logical Theorem Proving Planning Probabilistic Reasoning Machine Learning Solvers (eg MIP) Bioinformatics (Sequence Alignment, Protein Folding) Quant Finance (I read Wilmott’s books – interesting) Sooner or later, I’ll cover the above topics as well.

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• C#/.NET Little Wonders: The Concurrent Collections (1 of 3)

  - by James Michael Hare

  Once again we consider some of the lesser known classes and keywords of C#.  In the next few weeks, we will discuss the concurrent collections and how they have changed the face of concurrent programming. This week’s post will begin with a general introduction and discuss the ConcurrentStack<T> and ConcurrentQueue<T>.  Then in the following post we’ll discuss the ConcurrentDictionary<T> and ConcurrentBag<T>.  Finally, we shall close on the third post with a discussion of the BlockingCollection<T>. For more of the "Little Wonders" posts, see the index here. A brief history of collections In the beginning was the .NET 1.0 Framework.  And out of this framework emerged the System.Collections namespace, and it was good.  It contained all the basic things a growing programming language needs like the ArrayList and Hashtable collections.  The main problem, of course, with these original collections is that they held items of type object which means you had to be disciplined enough to use them correctly or you could end up with runtime errors if you got an object of a type you weren't expecting. Then came .NET 2.0 and generics and our world changed forever!  With generics the C# language finally got an equivalent of the very powerful C++ templates.  As such, the System.Collections.Generic was born and we got type-safe versions of all are favorite collections.  The List<T> succeeded the ArrayList and the Dictionary<TKey,TValue> succeeded the Hashtable and so on.  The new versions of the library were not only safer because they checked types at compile-time, in many cases they were more performant as well.  So much so that it's Microsoft's recommendation that the System.Collections original collections only be used for backwards compatibility. So we as developers came to know and love the generic collections and took them into our hearts and embraced them.  The problem is, thread safety in both the original collections and the generic collections can be problematic, for very different reasons. Now, if you are only doing single-threaded development you may not care – after all, no locking is required.  Even if you do have multiple threads, if a collection is “load-once, read-many” you don’t need to do anything to protect that container from multi-threaded access, as illustrated below: 1: public static class OrderTypeTranslator 2: { 3: // because this dictionary is loaded once before it is ever accessed, we don't need to synchronize 4: // multi-threaded read access 5: private static readonly Dictionary<string, char> _translator = new Dictionary<string, char> 6: { 7: {"New", 'N'}, 8: {"Update", 'U'}, 9: {"Cancel", 'X'} 10: }; 11:  12: // the only public interface into the dictionary is for reading, so inherently thread-safe 13: public static char? Translate(string orderType) 14: { 15: char charValue; 16: if (_translator.TryGetValue(orderType, out charValue)) 17: { 18: return charValue; 19: } 20:  21: return null; 22: } 23: } Unfortunately, most of our computer science problems cannot get by with just single-threaded applications or with multi-threading in a load-once manner.  Looking at  today's trends, it's clear to see that computers are not so much getting faster because of faster processor speeds -- we've nearly reached the limits we can push through with today's technologies -- but more because we're adding more cores to the boxes.  With this new hardware paradigm, it is even more important to use multi-threaded applications to take full advantage of parallel processing to achieve higher application speeds. So let's look at how to use collections in a thread-safe manner. Using historical collections in a concurrent fashion The early .NET collections (System.Collections) had a Synchronized() static method that could be used to wrap the early collections to make them completely thread-safe.  This paradigm was dropped in the generic collections (System.Collections.Generic) because having a synchronized wrapper resulted in atomic locks for all operations, which could prove overkill in many multithreading situations.  Thus the paradigm shifted to having the user of the collection specify their own locking, usually with an external object: 1: public class OrderAggregator 2: { 3: private static readonly Dictionary<string, List<Order>> _orders = new Dictionary<string, List<Order>>(); 4: private static readonly _orderLock = new object(); 5:  6: public void Add(string accountNumber, Order newOrder) 7: { 8: List<Order> ordersForAccount; 9:  10: // a complex operation like this should all be protected 11: lock (_orderLock) 12: { 13: if (!_orders.TryGetValue(accountNumber, out ordersForAccount)) 14: { 15: _orders.Add(accountNumber, ordersForAccount = new List<Order>()); 16: } 17:  18: ordersForAccount.Add(newOrder); 19: } 20: } 21: } Notice how we’re performing several operations on the dictionary under one lock.  With the Synchronized() static methods of the early collections, you wouldn’t be able to specify this level of locking (a more macro-level).  So in the generic collections, it was decided that if a user needed synchronization, they could implement their own locking scheme instead so that they could provide synchronization as needed. The need for better concurrent access to collections Here’s the problem: it’s relatively easy to write a collection that locks itself down completely for access, but anything more complex than that can be difficult and error-prone to write, and much less to make it perform efficiently!  For example, what if you have a Dictionary that has frequent reads but in-frequent updates?  Do you want to lock down the entire Dictionary for every access?  This would be overkill and would prevent concurrent reads.  In such cases you could use something like a ReaderWriterLockSlim which allows for multiple readers in a lock, and then once a writer grabs the lock it blocks all further readers until the writer is done (in a nutshell).  This is all very complex stuff to consider. Fortunately, this is where the Concurrent Collections come in.  The Parallel Computing Platform team at Microsoft went through great pains to determine how to make a set of concurrent collections that would have the best performance characteristics for general case multi-threaded use. Now, as in all things involving threading, you should always make sure you evaluate all your container options based on the particular usage scenario and the degree of parallelism you wish to acheive. This article should not be taken to understand that these collections are always supperior to the generic collections. Each fills a particular need for a particular situation. Understanding what each container is optimized for is key to the success of your application whether it be single-threaded or multi-threaded. General points to consider with the concurrent collections The MSDN points out that the concurrent collections all support the ICollection interface. However, since the collections are already synchronized, the IsSynchronized property always returns false, and SyncRoot always returns null.  Thus you should not attempt to use these properties for synchronization purposes. Note that since the concurrent collections also may have different operations than the traditional data structures you may be used to.  Now you may ask why they did this, but it was done out of necessity to keep operations safe and atomic.  For example, in order to do a Pop() on a stack you have to know the stack is non-empty, but between the time you check the stack’s IsEmpty property and then do the Pop() another thread may have come in and made the stack empty!  This is why some of the traditional operations have been changed to make them safe for concurrent use. In addition, some properties and methods in the concurrent collections achieve concurrency by creating a snapshot of the collection, which means that some operations that were traditionally O(1) may now be O(n) in the concurrent models.  I’ll try to point these out as we talk about each collection so you can be aware of any potential performance impacts.  Finally, all the concurrent containers are safe for enumeration even while being modified, but some of the containers support this in different ways (snapshot vs. dirty iteration).  Once again I’ll highlight how thread-safe enumeration works for each collection. ConcurrentStack<T>: The thread-safe LIFO container The ConcurrentStack<T> is the thread-safe counterpart to the System.Collections.Generic.Stack<T>, which as you may remember is your standard last-in-first-out container.  If you think of algorithms that favor stack usage (for example, depth-first searches of graphs and trees) then you can see how using a thread-safe stack would be of benefit. The ConcurrentStack<T> achieves thread-safe access by using System.Threading.Interlocked operations.  This means that the multi-threaded access to the stack requires no traditional locking and is very, very fast! For the most part, the ConcurrentStack<T> behaves like it’s Stack<T> counterpart with a few differences: Pop() was removed in favor of TryPop() Returns true if an item existed and was popped and false if empty. PushRange() and TryPopRange() were added Allows you to push multiple items and pop multiple items atomically. Count takes a snapshot of the stack and then counts the items. This means it is a O(n) operation, if you just want to check for an empty stack, call IsEmpty instead which is O(1). ToArray() and GetEnumerator() both also take snapshots. This means that iteration over a stack will give you a static view at the time of the call and will not reflect updates. Pushing on a ConcurrentStack<T> works just like you’d expect except for the aforementioned PushRange() method that was added to allow you to push a range of items concurrently. 1: var stack = new ConcurrentStack<string>(); 2:  3: // adding to stack is much the same as before 4: stack.Push("First"); 5:  6: // but you can also push multiple items in one atomic operation (no interleaves) 7: stack.PushRange(new [] { "Second", "Third", "Fourth" }); For looking at the top item of the stack (without removing it) the Peek() method has been removed in favor of a TryPeek().  This is because in order to do a peek the stack must be non-empty, but between the time you check for empty and the time you execute the peek the stack contents may have changed.  Thus the TryPeek() was created to be an atomic check for empty, and then peek if not empty: 1: // to look at top item of stack without removing it, can use TryPeek. 2: // Note that there is no Peek(), this is because you need to check for empty first. TryPeek does. 3: string item; 4: if (stack.TryPeek(out item)) 5: { 6: Console.WriteLine("Top item was " + item); 7: } 8: else 9: { 10: Console.WriteLine("Stack was empty."); 11: } Finally, to remove items from the stack, we have the TryPop() for single, and TryPopRange() for multiple items.  Just like the TryPeek(), these operations replace Pop() since we need to ensure atomically that the stack is non-empty before we pop from it: 1: // to remove items, use TryPop or TryPopRange to get multiple items atomically (no interleaves) 2: if (stack.TryPop(out item)) 3: { 4: Console.WriteLine("Popped " + item); 5: } 6:  7: // TryPopRange will only pop up to the number of spaces in the array, the actual number popped is returned. 8: var poppedItems = new string[2]; 9: int numPopped = stack.TryPopRange(poppedItems); 10:  11: foreach (var theItem in poppedItems.Take(numPopped)) 12: { 13: Console.WriteLine("Popped " + theItem); 14: } Finally, note that as stated before, GetEnumerator() and ToArray() gets a snapshot of the data at the time of the call.  That means if you are enumerating the stack you will get a snapshot of the stack at the time of the call.  This is illustrated below: 1: var stack = new ConcurrentStack<string>(); 2:  3: // adding to stack is much the same as before 4: stack.Push("First"); 5:  6: var results = stack.GetEnumerator(); 7:  8: // but you can also push multiple items in one atomic operation (no interleaves) 9: stack.PushRange(new [] { "Second", "Third", "Fourth" }); 10:  11: while(results.MoveNext()) 12: { 13: Console.WriteLine("Stack only has: " + results.Current); 14: } The only item that will be printed out in the above code is "First" because the snapshot was taken before the other items were added. This may sound like an issue, but it’s really for safety and is more correct.  You don’t want to enumerate a stack and have half a view of the stack before an update and half a view of the stack after an update, after all.  In addition, note that this is still thread-safe, whereas iterating through a non-concurrent collection while updating it in the old collections would cause an exception. ConcurrentQueue<T>: The thread-safe FIFO container The ConcurrentQueue<T> is the thread-safe counterpart of the System.Collections.Generic.Queue<T> class.  The concurrent queue uses an underlying list of small arrays and lock-free System.Threading.Interlocked operations on the head and tail arrays.  Once again, this allows us to do thread-safe operations without the need for heavy locks! The ConcurrentQueue<T> (like the ConcurrentStack<T>) has some departures from the non-concurrent counterpart.  Most notably: Dequeue() was removed in favor of TryDequeue(). Returns true if an item existed and was dequeued and false if empty. Count does not take a snapshot It subtracts the head and tail index to get the count.  This results overall in a O(1) complexity which is quite good.  It’s still recommended, however, that for empty checks you call IsEmpty instead of comparing Count to zero. ToArray() and GetEnumerator() both take snapshots. This means that iteration over a queue will give you a static view at the time of the call and will not reflect updates. The Enqueue() method on the ConcurrentQueue<T> works much the same as the generic Queue<T>: 1: var queue = new ConcurrentQueue<string>(); 2:  3: // adding to queue is much the same as before 4: queue.Enqueue("First"); 5: queue.Enqueue("Second"); 6: queue.Enqueue("Third"); For front item access, the TryPeek() method must be used to attempt to see the first item if the queue.  There is no Peek() method since, as you’ll remember, we can only peek on a non-empty queue, so we must have an atomic TryPeek() that checks for empty and then returns the first item if the queue is non-empty. 1: // to look at first item in queue without removing it, can use TryPeek. 2: // Note that there is no Peek(), this is because you need to check for empty first. TryPeek does. 3: string item; 4: if (queue.TryPeek(out item)) 5: { 6: Console.WriteLine("First item was " + item); 7: } 8: else 9: { 10: Console.WriteLine("Queue was empty."); 11: } Then, to remove items you use TryDequeue().  Once again this is for the same reason we have TryPeek() and not Peek(): 1: // to remove items, use TryDequeue. If queue is empty returns false. 2: if (queue.TryDequeue(out item)) 3: { 4: Console.WriteLine("Dequeued first item " + item); 5: } Just like the concurrent stack, the ConcurrentQueue<T> takes a snapshot when you call ToArray() or GetEnumerator() which means that subsequent updates to the queue will not be seen when you iterate over the results.  Thus once again the code below will only show the first item, since the other items were added after the snapshot. 1: var queue = new ConcurrentQueue<string>(); 2:  3: // adding to queue is much the same as before 4: queue.Enqueue("First"); 5:  6: var iterator = queue.GetEnumerator(); 7:  8: queue.Enqueue("Second"); 9: queue.Enqueue("Third"); 10:  11: // only shows First 12: while (iterator.MoveNext()) 13: { 14: Console.WriteLine("Dequeued item " + iterator.Current); 15: } Using collections concurrently You’ll notice in the examples above I stuck to using single-threaded examples so as to make them deterministic and the results obvious.  Of course, if we used these collections in a truly multi-threaded way the results would be less deterministic, but would still be thread-safe and with no locking on your part required! For example, say you have an order processor that takes an IEnumerable<Order> and handles each other in a multi-threaded fashion, then groups the responses together in a concurrent collection for aggregation.  This can be done easily with the TPL’s Parallel.ForEach(): 1: public static IEnumerable<OrderResult> ProcessOrders(IEnumerable<Order> orderList) 2: { 3: var proxy = new OrderProxy(); 4: var results = new ConcurrentQueue<OrderResult>(); 5:  6: // notice that we can process all these in parallel and put the results 7: // into our concurrent collection without needing any external locking! 8: Parallel.ForEach(orderList, 9: order => 10: { 11: var result = proxy.PlaceOrder(order); 12:  13: results.Enqueue(result); 14: }); 15:  16: return results; 17: } Summary Obviously, if you do not need multi-threaded safety, you don’t need to use these collections, but when you do need multi-threaded collections these are just the ticket! The plethora of features (I always think of the movie The Three Amigos when I say plethora) built into these containers and the amazing way they acheive thread-safe access in an efficient manner is wonderful to behold. Stay tuned next week where we’ll continue our discussion with the ConcurrentBag<T> and the ConcurrentDictionary<TKey,TValue>. For some excellent information on the performance of the concurrent collections and how they perform compared to a traditional brute-force locking strategy, see this wonderful whitepaper by the Microsoft Parallel Computing Platform team here.   Tweet Technorati Tags: C#,.NET,Concurrent Collections,Collections,Multi-Threading,Little Wonders,BlackRabbitCoder,James Michael Hare

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• Extending NerdDinner: Adding Geolocated Flair

  - by Jon Galloway

  NerdDinner is a website with the audacious goal of “Organizing the world’s nerds and helping them eat in packs.” Because nerds aren’t likely to socialize with others unless a website tells them to do it. Scott Hanselman showed off a lot of the cool features we’ve added to NerdDinner lately during his popular talk at MIX10, Beyond File | New Company: From Cheesy Sample to Social Platform. Did you miss it? Go ahead and watch it, I’ll wait. One of the features we wanted to add was flair. You know about flair, right? It’s a way to let folks who like your site show it off in their own site. For example, here’s my StackOverflow flair: Great! So how could we add some of this flair stuff to NerdDinner? What do we want to show? If we’re going to encourage our users to give up a bit of their beautiful website to show off a bit of ours, we need to think about what they’ll want to show. For instance, my StackOverflow flair is all about me, not StackOverflow. So how will this apply to NerdDinner? Since NerdDinner is all about organizing local dinners, in order for the flair to be useful it needs to make sense for the person viewing the web page. If someone visits from Egypt visits my blog, they should see information about NerdDinners in Egypt. That’s geolocation – localizing site content based on where the browser’s sitting, and it makes sense for flair as well as entire websites. So we’ll set up a simple little callout that prompts them to host a dinner in their area: Hopefully our flair works and there is a dinner near your viewers, so they’ll see another view which lists upcoming dinners near them: The Geolocation Part Generally website geolocation is done by mapping the requestor’s IP address to a geographic area. It’s not an exact science, but I’ve always found it to be pretty accurate. There are (at least) three ways to handle it: You pay somebody like MaxMind for a database (with regular updates) that sits on your server, and you use their API to do lookups. I used this on a pretty big project a few years ago and it worked well. You use HTML 5 Geolocation API or Google Gears or some other browser based solution. I think those are cool (I use Google Gears a lot), but they’re both in flux right now and I don’t think either has a wide enough of an install base yet to rely on them. You might want to, but I’ve heard you do all kinds of crazy stuff, and sometimes it gets you in trouble. I don’t mean talk out of line, but we all laugh behind your back a bit. But, hey, it’s up to you. It’s your flair or whatever. There are some free webservices out there that will take an IP address and give you location information. Easy, and works for everyone. That’s what we’re doing. I looked at a few different services and settled on IPInfoDB. It’s free, has a great API, and even returns JSON, which is handy for Javascript use. The IP query is pretty simple. We hit a URL like this: http://ipinfodb.com/ip_query.php?ip=74.125.45.100&timezone=false … and we get an XML response back like this… <?xml version="1.0" encoding="UTF-8"?> <Response> <Ip>74.125.45.100</Ip> <Status>OK</Status> <CountryCode>US</CountryCode> <CountryName>United States</CountryName> <RegionCode>06</RegionCode> <RegionName>California</RegionName> <City>Mountain View</City> <ZipPostalCode>94043</ZipPostalCode> <Latitude>37.4192</Latitude> <Longitude>-122.057</Longitude> </Response> So we’ll build some data transfer classes to hold the location information, like this: public class LocationInfo { public string Country { get; set; } public string RegionName { get; set; } public string City { get; set; } public string ZipPostalCode { get; set; } public LatLong Position { get; set; } } public class LatLong { public float Lat { get; set; } public float Long { get; set; } } And now hitting the service is pretty simple: public static LocationInfo HostIpToPlaceName(string ip) { string url = "http://ipinfodb.com/ip_query.php?ip={0}&timezone=false"; url = String.Format(url, ip); var result = XDocument.Load(url); var location = (from x in result.Descendants("Response") select new LocationInfo { City = (string)x.Element("City"), RegionName = (string)x.Element("RegionName"), Country = (string)x.Element("CountryName"), ZipPostalCode = (string)x.Element("CountryName"), Position = new LatLong { Lat = (float)x.Element("Latitude"), Long = (float)x.Element("Longitude") } }).First(); return location; } Getting The User’s IP Okay, but first we need the end user’s IP, and you’d think it would be as simple as reading the value from HttpContext: HttpContext.Current.Request.UserHostAddress But you’d be wrong. Sorry. UserHostAddress just wraps HttpContext.Current.Request.ServerVariables["REMOTE_ADDR"], but that doesn’t get you the IP for users behind a proxy. That’s in another header, “HTTP_X_FORWARDED_FOR". So you can either hit a wrapper and then check a header, or just check two headers. I went for uniformity: string SourceIP = string.IsNullOrEmpty(Request.ServerVariables["HTTP_X_FORWARDED_FOR"]) ? Request.ServerVariables["REMOTE_ADDR"] : Request.ServerVariables["HTTP_X_FORWARDED_FOR"]; We’re almost set to wrap this up, but first let’s talk about our views. Yes, views, because we’ll have two. Selecting the View We wanted to make it easy for people to include the flair in their sites, so we looked around at how other people were doing this. The StackOverflow folks have a pretty good flair system, which allows you to include the flair in your site as either an IFRAME reference or a Javascript include. We’ll do both. We have a ServicesController to handle use of the site information outside of NerdDinner.com, so this fits in pretty well there. We’ll be displaying the same information for both HTML and Javascript flair, so we can use one Flair controller action which will return a different view depending on the requested format. Here’s our general flow for our controller action: Get the user’s IP Translate it to a location Grab the top three upcoming dinners that are near that location Select the view based on the format (defaulted to “html”) Return a FlairViewModel which contains the list of dinners and the location information public ActionResult Flair(string format = "html") { string SourceIP = string.IsNullOrEmpty( Request.ServerVariables["HTTP_X_FORWARDED_FOR"]) ? Request.ServerVariables["REMOTE_ADDR"] : Request.ServerVariables["HTTP_X_FORWARDED_FOR"]; var location = GeolocationService.HostIpToPlaceName(SourceIP); var dinners = dinnerRepository. FindByLocation(location.Position.Lat, location.Position.Long). OrderByDescending(p => p.EventDate).Take(3); // Select the view we'll return. // Using a switch because we'll add in JSON and other formats later. string view; switch (format.ToLower()) { case "javascript": view = "JavascriptFlair"; break; default: view = "Flair"; break; } return View( view, new FlairViewModel { Dinners = dinners.ToList(), LocationName = string.IsNullOrEmpty(location.City) ? "you" : String.Format("{0}, {1}", location.City, location.RegionName) } ); } Note: I’m not in love with the logic here, but it seems like overkill to extract the switch statement away when we’ll probably just have two or three views. What do you think? The HTML View The HTML version of the view is pretty simple – the only thing of any real interest here is the use of an extension method to truncate strings that are would cause the titles to wrap. public static string Truncate(this string s, int maxLength) { if (string.IsNullOrEmpty(s) || maxLength <= 0) return string.Empty; else if (s.Length > maxLength) return s.Substring(0, maxLength) + "..."; else return s; }   So here’s how the HTML view ends up looking: <%@ Page Title="" Language="C#" Inherits="System.Web.Mvc.ViewPage<FlairViewModel>" %> <%@ Import Namespace="NerdDinner.Helpers" %> <%@ Import Namespace="NerdDinner.Models" %> <!DOCTYPE html PUBLIC "-//W3C//DTD XHTML 1.0 Strict//EN" "http://www.w3.org/TR/xhtml1/DTD/xhtml1-strict.dtd"> <html xmlns="http://www.w3.org/1999/xhtml"> <head> <title>Nerd Dinner</title> <link href="/Content/Flair.css" rel="stylesheet" type="text/css" /> </head> <body> <div id="nd-wrapper"> <h2 id="nd-header">NerdDinner.com</h2> <div id="nd-outer"> <% if (Model.Dinners.Count == 0) { %> <div id="nd-bummer"> Looks like there's no Nerd Dinners near <%:Model.LocationName %> in the near future. Why not <a target="_blank" href="http://www.nerddinner.com/Dinners/Create">host one</a>?</div> <% } else { %> <h3> Dinners Near You</h3> <ul> <% foreach (var item in Model.Dinners) { %> <li> <%: Html.ActionLink(String.Format("{0} with {1} on {2}", item.Title.Truncate(20), item.HostedBy, item.EventDate.ToShortDateString()), "Details", "Dinners", new { id = item.DinnerID }, new { target = "_blank" })%></li> <% } %> </ul> <% } %> <div id="nd-footer"> More dinners and fun at <a target="_blank" href="http://nrddnr.com">http://nrddnr.com</a></div> </div> </div> </body> </html> You’d include this in a page using an IFRAME, like this: <IFRAME height=230 marginHeight=0 src="http://nerddinner.com/services/flair" frameBorder=0 width=160 marginWidth=0 scrolling=no></IFRAME> The Javascript view The Javascript flair is written so you can include it in a webpage with a simple script include, like this: <script type="text/javascript" src="http://nerddinner.com/services/flair?format=javascript"></script> The goal of this view is very similar to the HTML embed view, with a few exceptions: We’re creating a script element and adding it to the head of the document, which will then document.write out the content. Note that you have to consider if your users will actually have a <head> element in their documents, but for website flair use cases I think that’s a safe bet. Since the content is being added to the existing page rather than shown in an IFRAME, all links need to be absolute. That means we can’t use Html.ActionLink, since it generates relative routes. We need to escape everything since it’s being written out as strings. We need to set the content type to application/x-javascript. The easiest way to do that is to use the <%@ Page ContentType%> directive. <%@ Page Language="C#" Inherits="System.Web.Mvc.ViewPage<NerdDinner.Models.FlairViewModel>" ContentType="application/x-javascript" %> <%@ Import Namespace="NerdDinner.Helpers" %> <%@ Import Namespace="NerdDinner.Models" %> document.write('<script>var link = document.createElement(\"link\");link.href = \"http://nerddinner.com/content/Flair.css\";link.rel = \"stylesheet\";link.type = \"text/css\";var head = document.getElementsByTagName(\"head\")[0];head.appendChild(link);</script>'); document.write('<div id=\"nd-wrapper\"><h2 id=\"nd-header\">NerdDinner.com</h2><div id=\"nd-outer\">'); <% if (Model.Dinners.Count == 0) { %> document.write('<div id=\"nd-bummer\">Looks like there\'s no Nerd Dinners near <%:Model.LocationName %> in the near future. Why not <a target=\"_blank\" href=\"http://www.nerddinner.com/Dinners/Create\">host one</a>?</div>'); <% } else { %> document.write('<h3> Dinners Near You</h3><ul>'); <% foreach (var item in Model.Dinners) { %> document.write('<li><a target=\"_blank\" href=\"http://nrddnr.com/<%: item.DinnerID %>\"><%: item.Title.Truncate(20) %> with <%: item.HostedBy %> on <%: item.EventDate.ToShortDateString() %></a></li>'); <% } %> document.write('</ul>'); <% } %> document.write('<div id=\"nd-footer\"> More dinners and fun at <a target=\"_blank\" href=\"http://nrddnr.com\">http://nrddnr.com</a></div></div></div>'); Getting IP’s for Testing There are a variety of online services that will translate a location to an IP, which were handy for testing these out. I found http://www.itouchmap.com/latlong.html to be most useful, but I’m open to suggestions if you know of something better. Next steps I think the next step here is to minimize load – you know, in case people start actually using this flair. There are two places to think about – the NerdDinner.com servers, and the services we’re using for Geolocation. I usually think about caching as a first attack on server load, but that’s less helpful here since every user will have a different IP. Instead, I’d look at taking advantage of Asynchronous Controller Actions, a cool new feature in ASP.NET MVC 2. Async Actions let you call a potentially long-running webservice without tying up a thread on the server while waiting for the response. There’s some good info on that in the MSDN documentation, and Dino Esposito wrote a great article on Asynchronous ASP.NET Pages in the April 2010 issue of MSDN Magazine. But let’s think of the children, shall we? What about ipinfodb.com? Well, they don’t have specific daily limits, but they do throttle you if you put a lot of traffic on them. From their FAQ: We do not have a specific daily limit but queries that are at a rate faster than 2 per second will be put in "queue". If you stay below 2 queries/second everything will be normal. If you go over the limit, you will still get an answer for all queries but they will be slowed down to about 1 per second. This should not affect most users but for high volume websites, you can either use our IP database on your server or we can whitelist your IP for 5$/month (simply use the donate form and leave a comment with your server IP). Good programming practices such as not querying our API for all page views (you can store the data in a cookie or a database) will also help not reaching the limit. So the first step there is to save the geolocalization information in a time-limited cookie, which will allow us to look up the local dinners immediately without having to hit the geolocation service.

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• Using Stub Objects

  - by user9154181

  Having told the long and winding tale of where stub objects came from and how we use them to build Solaris, I'd like to focus now on the the nuts and bolts of building and using them. The following new features were added to the Solaris link-editor (ld) to support the production and use of stub objects: -z stub This new command line option informs ld that it is to build a stub object rather than a normal object. In this mode, it accepts the same command line arguments as usual, but will quietly ignore any objects and sharable object dependencies. STUB_OBJECT Mapfile Directive In order to build a stub version of an object, its mapfile must specify the STUB_OBJECT directive. When producing a non-stub object, the presence of STUB_OBJECT causes the link-editor to perform extra validation to ensure that the stub and non-stub objects will be compatible. ASSERT Mapfile Directive All data symbols exported from the object must have an ASSERT symbol directive in the mapfile that declares them as data and supplies the size, binding, bss attributes, and symbol aliasing details. When building the stub objects, the information in these ASSERT directives is used to create the data symbols. When building the real object, these ASSERT directives will ensure that the real object matches the linking interface presented by the stub. Although ASSERT was added to the link-editor in order to support stub objects, they are a general purpose feature that can be used independently of stub objects. For instance you might choose to use an ASSERT directive if you have a symbol that must have a specific address in order for the object to operate properly and you want to automatically ensure that this will always be the case. The material presented here is derived from a document I originally wrote during the development effort, which had the dual goals of providing supplemental materials for the stub object PSARC case, and as a set of edits that were eventually applied to the Oracle Solaris Linker and Libraries Manual (LLM). The Solaris 11 LLM contains this information in a more polished form. Stub Objects A stub object is a shared object, built entirely from mapfiles, that supplies the same linking interface as the real object, while containing no code or data. Stub objects cannot be used at runtime. However, an application can be built against a stub object, where the stub object provides the real object name to be used at runtime, and then use the real object at runtime. When building a stub object, the link-editor ignores any object or library files specified on the command line, and these files need not exist in order to build a stub. Since the compilation step can be omitted, and because the link-editor has relatively little work to do, stub objects can be built very quickly. Stub objects can be used to solve a variety of build problems: Speed Modern machines, using a version of make with the ability to parallelize operations, are capable of compiling and linking many objects simultaneously, and doing so offers significant speedups. However, it is typical that a given object will depend on other objects, and that there will be a core set of objects that nearly everything else depends on. It is necessary to impose an ordering that builds each object before any other object that requires it. This ordering creates bottlenecks that reduce the amount of parallelization that is possible and limits the overall speed at which the code can be built. Complexity/Correctness In a large body of code, there can be a large number of dependencies between the various objects. The makefiles or other build descriptions for these objects can become very complex and difficult to understand or maintain. The dependencies can change as the system evolves. This can cause a given set of makefiles to become slightly incorrect over time, leading to race conditions and mysterious rare build failures. Dependency Cycles It might be desirable to organize code as cooperating shared objects, each of which draw on the resources provided by the other. Such cycles cannot be supported in an environment where objects must be built before the objects that use them, even though the runtime linker is fully capable of loading and using such objects if they could be built. Stub shared objects offer an alternative method for building code that sidesteps the above issues. Stub objects can be quickly built for all the shared objects produced by the build. Then, all the real shared objects and executables can be built in parallel, in any order, using the stub objects to stand in for the real objects at link-time. Afterwards, the executables and real shared objects are kept, and the stub shared objects are discarded. Stub objects are built from a mapfile, which must satisfy the following requirements. The mapfile must specify the STUB_OBJECT directive. This directive informs the link-editor that the object can be built as a stub object, and as such causes the link-editor to perform validation and sanity checking intended to guarantee that an object and its stub will always provide identical linking interfaces. All function and data symbols that make up the external interface to the object must be explicitly listed in the mapfile. The mapfile must use symbol scope reduction ('*'), to remove any symbols not explicitly listed from the external interface. All global data exported from the object must have an ASSERT symbol attribute in the mapfile to specify the symbol type, size, and bss attributes. In the case where there are multiple symbols that reference the same data, the ASSERT for one of these symbols must specify the TYPE and SIZE attributes, while the others must use the ALIAS attribute to reference this primary symbol. Given such a mapfile, the stub and real versions of the shared object can be built using the same command line for each, adding the '-z stub' option to the link for the stub object, and omiting the option from the link for the real object. To demonstrate these ideas, the following code implements a shared object named idx5, which exports data from a 5 element array of integers, with each element initialized to contain its zero-based array index. This data is available as a global array, via an alternative alias data symbol with weak binding, and via a functional interface. % cat idx5.c int _idx5[5] = { 0, 1, 2, 3, 4 }; #pragma weak idx5 = _idx5 int idx5_func(int index) { if ((index 4)) return (-1); return (_idx5[index]); } A mapfile is required to describe the interface provided by this shared object. % cat mapfile $mapfile_version 2 STUB_OBJECT; SYMBOL_SCOPE { _idx5 { ASSERT { TYPE=data; SIZE=4[5] }; }; idx5 { ASSERT { BINDING=weak; ALIAS=_idx5 }; }; idx5_func; local: *; }; The following main program is used to print all the index values available from the idx5 shared object. % cat main.c #include <stdio.h> extern int _idx5[5], idx5[5], idx5_func(int); int main(int argc, char **argv) { int i; for (i = 0; i The following commands create a stub version of this shared object in a subdirectory named stublib. elfdump is used to verify that the resulting object is a stub. The command used to build the stub differs from that of the real object only in the addition of the -z stub option, and the use of a different output file name. This demonstrates the ease with which stub generation can be added to an existing makefile. % cc -Kpic -G -M mapfile -h libidx5.so.1 idx5.c -o stublib/libidx5.so.1 -zstub % ln -s libidx5.so.1 stublib/libidx5.so % elfdump -d stublib/libidx5.so | grep STUB [11] FLAGS_1 0x4000000 [ STUB ] The main program can now be built, using the stub object to stand in for the real shared object, and setting a runpath that will find the real object at runtime. However, as we have not yet built the real object, this program cannot yet be run. Attempts to cause the system to load the stub object are rejected, as the runtime linker knows that stub objects lack the actual code and data found in the real object, and cannot execute. % cc main.c -L stublib -R '$ORIGIN/lib' -lidx5 -lc % ./a.out ld.so.1: a.out: fatal: libidx5.so.1: open failed: No such file or directory Killed % LD_PRELOAD=stublib/libidx5.so.1 ./a.out ld.so.1: a.out: fatal: stublib/libidx5.so.1: stub shared object cannot be used at runtime Killed We build the real object using the same command as we used to build the stub, omitting the -z stub option, and writing the results to a different file. % cc -Kpic -G -M mapfile -h libidx5.so.1 idx5.c -o lib/libidx5.so.1 Once the real object has been built in the lib subdirectory, the program can be run. % ./a.out [0] 0 0 0 [1] 1 1 1 [2] 2 2 2 [3] 3 3 3 [4] 4 4 4 Mapfile Changes The version 2 mapfile syntax was extended in a number of places to accommodate stub objects. Conditional Input The version 2 mapfile syntax has the ability conditionalize mapfile input using the $if control directive. As you might imagine, these directives are used frequently with ASSERT directives for data, because a given data symbol will frequently have a different size in 32 or 64-bit code, or on differing hardware such as x86 versus sparc. The link-editor maintains an internal table of names that can be used in the logical expressions evaluated by $if and $elif. At startup, this table is initialized with items that describe the class of object (_ELF32 or _ELF64) and the type of the target machine (_sparc or _x86). We found that there were a small number of cases in the Solaris code base in which we needed to know what kind of object we were producing, so we added the following new predefined items in order to address that need: NameMeaning ...... _ET_DYNshared object _ET_EXECexecutable object _ET_RELrelocatable object ...... STUB_OBJECT Directive The new STUB_OBJECT directive informs the link-editor that the object described by the mapfile can be built as a stub object. STUB_OBJECT; A stub shared object is built entirely from the information in the mapfiles supplied on the command line. When the -z stub option is specified to build a stub object, the presence of the STUB_OBJECT directive in a mapfile is required, and the link-editor uses the information in symbol ASSERT attributes to create global symbols that match those of the real object. When the real object is built, the presence of STUB_OBJECT causes the link-editor to verify that the mapfiles accurately describe the real object interface, and that a stub object built from them will provide the same linking interface as the real object it represents. All function and data symbols that make up the external interface to the object must be explicitly listed in the mapfile. The mapfile must use symbol scope reduction ('*'), to remove any symbols not explicitly listed from the external interface. All global data in the object is required to have an ASSERT attribute that specifies the symbol type and size. If the ASSERT BIND attribute is not present, the link-editor provides a default assertion that the symbol must be GLOBAL. If the ASSERT SH_ATTR attribute is not present, or does not specify that the section is one of BITS or NOBITS, the link-editor provides a default assertion that the associated section is BITS. All data symbols that describe the same address and size are required to have ASSERT ALIAS attributes specified in the mapfile. If aliased symbols are discovered that do not have an ASSERT ALIAS specified, the link fails and no object is produced. These rules ensure that the mapfiles contain a description of the real shared object's linking interface that is sufficient to produce a stub object with a completely compatible linking interface. SYMBOL_SCOPE/SYMBOL_VERSION ASSERT Attribute The SYMBOL_SCOPE and SYMBOL_VERSION mapfile directives were extended with a symbol attribute named ASSERT. The syntax for the ASSERT attribute is as follows: ASSERT { ALIAS = symbol_name; BINDING = symbol_binding; TYPE = symbol_type; SH_ATTR = section_attributes; SIZE = size_value; SIZE = size_value[count]; }; The ASSERT attribute is used to specify the expected characteristics of the symbol. The link-editor compares the symbol characteristics that result from the link to those given by ASSERT attributes. If the real and asserted attributes do not agree, a fatal error is issued and the output object is not created. In normal use, the link editor evaluates the ASSERT attribute when present, but does not require them, or provide default values for them. The presence of the STUB_OBJECT directive in a mapfile alters the interpretation of ASSERT to require them under some circumstances, and to supply default assertions if explicit ones are not present. See the definition of the STUB_OBJECT Directive for the details. When the -z stub command line option is specified to build a stub object, the information provided by ASSERT attributes is used to define the attributes of the global symbols provided by the object. ASSERT accepts the following: ALIAS Name of a previously defined symbol that this symbol is an alias for. An alias symbol has the same type, value, and size as the main symbol. The ALIAS attribute is mutually exclusive to the TYPE, SIZE, and SH_ATTR attributes, and cannot be used with them. When ALIAS is specified, the type, size, and section attributes are obtained from the alias symbol. BIND Specifies an ELF symbol binding, which can be any of the STB_ constants defined in <sys/elf.h>, with the STB_ prefix removed (e.g. GLOBAL, WEAK). TYPE Specifies an ELF symbol type, which can be any of the STT_ constants defined in <sys/elf.h>, with the STT_ prefix removed (e.g. OBJECT, COMMON, FUNC). In addition, for compatibility with other mapfile usage, FUNCTION and DATA can be specified, for STT_FUNC and STT_OBJECT, respectively. TYPE is mutually exclusive to ALIAS, and cannot be used in conjunction with it. SH_ATTR Specifies attributes of the section associated with the symbol. The section_attributes that can be specified are given in the following table: Section AttributeMeaning BITSSection is not of type SHT_NOBITS NOBITSSection is of type SHT_NOBITS SH_ATTR is mutually exclusive to ALIAS, and cannot be used in conjunction with it. SIZE Specifies the expected symbol size. SIZE is mutually exclusive to ALIAS, and cannot be used in conjunction with it. The syntax for the size_value argument is as described in the discussion of the SIZE attribute below. SIZE The SIZE symbol attribute existed before support for stub objects was introduced. It is used to set the size attribute of a given symbol. This attribute results in the creation of a symbol definition. Prior to the introduction of the ASSERT SIZE attribute, the value of a SIZE attribute was always numeric. While attempting to apply ASSERT SIZE to the objects in the Solaris ON consolidation, I found that many data symbols have a size based on the natural machine wordsize for the class of object being produced. Variables declared as long, or as a pointer, will be 4 bytes in size in a 32-bit object, and 8 bytes in a 64-bit object. Initially, I employed the conditional $if directive to handle these cases as follows: $if _ELF32 foo { ASSERT { TYPE=data; SIZE=4 } }; bar { ASSERT { TYPE=data; SIZE=20 } }; $elif _ELF64 foo { ASSERT { TYPE=data; SIZE=8 } }; bar { ASSERT { TYPE=data; SIZE=40 } }; $else $error UNKNOWN ELFCLASS $endif I found that the situation occurs frequently enough that this is cumbersome. To simplify this case, I introduced the idea of the addrsize symbolic name, and of a repeat count, which together make it simple to specify machine word scalar or array symbols. Both the SIZE, and ASSERT SIZE attributes support this syntax: The size_value argument can be a numeric value, or it can be the symbolic name addrsize. addrsize represents the size of a machine word capable of holding a memory address. The link-editor substitutes the value 4 for addrsize when building 32-bit objects, and the value 8 when building 64-bit objects. addrsize is useful for representing the size of pointer variables and C variables of type long, as it automatically adjusts for 32 and 64-bit objects without requiring the use of conditional input. The size_value argument can be optionally suffixed with a count value, enclosed in square brackets. If count is present, size_value and count are multiplied together to obtain the final size value. Using this feature, the example above can be written more naturally as: foo { ASSERT { TYPE=data; SIZE=addrsize } }; bar { ASSERT { TYPE=data; SIZE=addrsize[5] } }; Exported Global Data Is Still A Bad Idea As you can see, the additional plumbing added to the Solaris link-editor to support stub objects is minimal. Furthermore, about 90% of that plumbing is dedicated to handling global data. We have long advised against global data exported from shared objects. There are many ways in which global data does not fit well with dynamic linking. Stub objects simply provide one more reason to avoid this practice. It is always better to export all data via a functional interface. You should always hide your data, and make it available to your users via a function that they can call to acquire the address of the data item. However, If you do have to support global data for a stub, perhaps because you are working with an already existing object, it is still easilily done, as shown above. Oracle does not like us to discuss hypothetical new features that don't exist in shipping product, so I'll end this section with a speculation. It might be possible to do more in this area to ease the difficulty of dealing with objects that have global data that the users of the library don't need. Perhaps someday... Conclusions It is easy to create stub objects for most objects. If your library only exports function symbols, all you have to do to build a faithful stub object is to add STUB_OBJECT; and then to use the same link command you're currently using, with the addition of the -z stub option. Happy Stubbing!

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• Microsoft Introduces WebMatrix

  - by Rick Strahl

  originally published in CoDe Magazine Editorial Microsoft recently released the first CTP of a new development environment called WebMatrix, which along with some of its supporting technologies are squarely aimed at making the Microsoft Web Platform more approachable for first-time developers and hobbyists. But in the process, it also provides some updated technologies that can make life easier for existing .NET developers. Let’s face it: ASP.NET development isn’t exactly trivial unless you already have a fair bit of familiarity with sophisticated development practices. Stick a non-developer in front of Visual Studio .NET or even the Visual Web Developer Express edition and it’s not likely that the person in front of the screen will be very productive or feel inspired. Yet other technologies like PHP and even classic ASP did provide the ability for non-developers and hobbyists to become reasonably proficient in creating basic web content quickly and efficiently. WebMatrix appears to be Microsoft’s attempt to bring back some of that simplicity with a number of technologies and tools. The key is to provide a friendly and fully self-contained development environment that provides all the tools needed to build an application in one place, as well as tools that allow publishing of content and databases easily to the web server. WebMatrix is made up of several components and technologies: IIS Developer Express IIS Developer Express is a new, self-contained development web server that is fully compatible with IIS 7.5 and based on the same codebase that IIS 7.5 uses. This new development server replaces the much less compatible Cassini web server that’s been used in Visual Studio and the Express editions. IIS Express addresses a few shortcomings of the Cassini server such as the inability to serve custom ISAPI extensions (i.e., things like PHP or ASP classic for example), as well as not supporting advanced authentication. IIS Developer Express provides most of the IIS 7.5 feature set providing much better compatibility between development and live deployment scenarios. SQL Server Compact 4.0 Database access is a key component for most web-driven applications, but on the Microsoft stack this has mostly meant you have to use SQL Server or SQL Server Express. SQL Server Compact is not new-it’s been around for a few years, but it’s been severely hobbled in the past by terrible tool support and the inability to support more than a single connection in Microsoft’s attempt to avoid losing SQL Server licensing. The new release of SQL Server Compact 4.0 supports multiple connections and you can run it in ASP.NET web applications simply by installing an assembly into the bin folder of the web application. In effect, you don’t have to install a special system configuration to run SQL Compact as it is a drop-in database engine: Copy the small assembly into your BIN folder (or from the GAC if installed fully), create a connection string against a local file-based database file, and then start firing SQL requests. Additionally WebMatrix includes nice tools to edit the database tables and files, along with tools to easily upsize (and hopefully downsize in the future) to full SQL Server. This is a big win, pending compatibility and performance limits. In my simple testing the data engine performed well enough for small data sets. This is not only useful for web applications, but also for desktop applications for which a fully installed SQL engine like SQL Server would be overkill. Having a local data store in those applications that can potentially be accessed by multiple users is a welcome feature. ASP.NET Razor View Engine What? Yet another native ASP.NET view engine? We already have Web Forms and various different flavors of using that view engine with Web Forms and MVC. Do we really need another? Microsoft thinks so, and Razor is an implementation of a lightweight, script-only view engine. Unlike the Web Forms view engine, Razor works only with inline code, snippets, and markup; therefore, it is more in line with current thinking of what a view engine should represent. There’s no support for a “page model” or any of the other Web Forms features of the full-page framework, but just a lightweight scripting engine that works with plain markup plus embedded expressions and code. The markup syntax for Razor is geared for minimal typing, plus some progressive detection of where a script block/expression starts and ends. This results in a much leaner syntax than the typical ASP.NET Web Forms alligator (<% %>) tags. Razor uses the @ sign plus standard C# (or Visual Basic) block syntax to delineate code snippets and expressions. Here’s a very simple example of what Razor markup looks like along with some comment annotations: <!DOCTYPE html> <html>     <head>         <title></title>     </head>     <body>     <h1>Razor Test</h1>          <!-- simple expressions -->     @DateTime.Now     <hr />     <!-- method expressions -->     @DateTime.Now.ToString("T")          <!-- code blocks -->     @{         List<string> names = new List<string>();         names.Add("Rick");         names.Add("Markus");         names.Add("Claudio");         names.Add("Kevin");     }          <!-- structured block statements -->     <ul>     @foreach(string name in names){             <li>@name</li>     }     </ul>           <!-- Conditional code -->        @if(true) {                        <!-- Literal Text embedding in code -->        <text>         true        </text>;    }    else    {        <!-- Literal Text embedding in code -->       <text>       false       </text>;    }    </body> </html> Like the Web Forms view engine, Razor parses pages into code, and then executes that run-time compiled code. Effectively a “page” becomes a code file with markup becoming literal text written into the Response stream, code snippets becoming raw code, and expressions being written out with Response.Write(). The code generated from Razor doesn’t look much different from similar Web Forms code that only uses script tags; so although the syntax may look different, the operational model is fairly similar to the Web Forms engine minus the overhead of the large Page object model. However, there are differences: -Razor pages are based on a new base class, Microsoft.WebPages.WebPage, which is hosted in the Microsoft.WebPages assembly that houses all the Razor engine parsing and processing logic. Browsing through the assembly (in the generated ASP.NET Temporary Files folder or GAC) will give you a good idea of the functionality that Razor provides. If you look closely, a lot of the feature set matches ASP.NET MVC’s view implementation as well as many of the helper classes found in MVC. It’s not hard to guess the motivation for this sort of view engine: For beginning developers the simple markup syntax is easier to work with, although you obviously still need to have some understanding of the .NET Framework in order to create dynamic content. The syntax is easier to read and grok and much shorter to type than ASP.NET alligator tags (<% %>) and also easier to understand aesthetically what’s happening in the markup code. Razor also is a better fit for Microsoft’s vision of ASP.NET MVC: It’s a new view engine without the baggage of Web Forms attached to it. The engine is more lightweight since it doesn’t carry all the features and object model of Web Forms with it and it can be instantiated directly outside of the HTTP environment, which has been rather tricky to do for the Web Forms view engine. Having a standalone script parser is a huge win for other applications as well – it makes it much easier to create script or meta driven output generators for many types of applications from code/screen generators, to simple form letters to data merging applications with user customizability. For me personally this is very useful side effect and who knows maybe Microsoft will actually standardize they’re scripting engines (die T4 die!) on this engine. Razor also better fits the “view-based” approach where the view is supposed to be mostly a visual representation that doesn’t hold much, if any, code. While you can still use code, the code you do write has to be self-contained. Overall I wouldn’t be surprised if Razor will become the new standard view engine for MVC in the future – and in fact there have been announcements recently that Razor will become the default script engine in ASP.NET MVC 3.0. Razor can also be used in existing Web Forms and MVC applications, although that’s not working currently unless you manually configure the script mappings and add the appropriate assemblies. It’s possible to do it, but it’s probably better to wait until Microsoft releases official support for Razor scripts in Visual Studio. Once that happens, you can simply drop .cshtml and .vbhtml pages into an existing ASP.NET project and they will work side by side with classic ASP.NET pages. WebMatrix Development Environment To tie all of these three technologies together, Microsoft is shipping WebMatrix with an integrated development environment. An integrated gallery manager makes it easy to download and load existing projects, and then extend them with custom functionality. It seems to be a prominent goal to provide community-oriented content that can act as a starting point, be it via a custom templates or a complete standard application. The IDE includes a project manager that works with a single project and provides an integrated IDE/editor for editing the .cshtml and .vbhtml pages. A run button allows you to quickly run pages in the project manager in a variety of browsers. There’s no debugging support for code at this time. Note that Razor pages don’t require explicit compilation, so making a change, saving, and then refreshing your page in the browser is all that’s needed to see changes while testing an application locally. It’s essentially using the auto-compiling Web Project that was introduced with .NET 2.0. All code is compiled during run time into dynamically created assemblies in the ASP.NET temp folder. WebMatrix also has PHP Editing support with syntax highlighting. You can load various PHP-based applications from the WebMatrix Web Gallery directly into the IDE. Most of the Web Gallery applications are ready to install and run without further configuration, with Wizards taking you through installation of tools, dependencies, and configuration of the database as needed. WebMatrix leverages the Web Platform installer to pull the pieces down from websites in a tight integration of tools that worked nicely for the four or five applications I tried this out on. Click a couple of check boxes and fill in a few simple configuration options and you end up with a running application that’s ready to be customized. Nice! You can easily deploy completed applications via WebDeploy (to an IIS server) or FTP directly from within the development environment. The deploy tool also can handle automatically uploading and installing the database and all related assemblies required, making deployment a simple one-click install step. Simplified Database Access The IDE contains a database editor that can edit SQL Compact and SQL Server databases. There is also a Database helper class that facilitates database access by providing easy-to-use, high-level query execution and iteration methods: @{       var db = Database.OpenFile("FirstApp.sdf");     string sql = "select * from customers where Id > @0"; } <ul> @foreach(var row in db.Query(sql,1)){         <li>@row.FirstName @row.LastName</li> } </ul> The query function takes a SQL statement plus any number of positional (@0,@1 etc.) SQL parameters by simple values. The result is returned as a collection of rows which in turn have a row object with dynamic properties for each of the columns giving easy (though untyped) access to each of the fields. Likewise Execute and ExecuteNonQuery allow execution of more complex queries using similar parameter passing schemes. Note these queries use string-based queries rather than LINQ or Entity Framework’s strongly typed LINQ queries. While this may seem like a step back, it’s also in line with the expectations of non .NET script developers who are quite used to writing and using SQL strings in code rather than using OR/M frameworks. The only question is why was something not included from the beginning in .NET and Microsoft made developers build custom implementations of these basic building blocks. The implementation looks a lot like a DataTable-style data access mechanism, but to be fair, this is a common approach in scripting languages. This type of syntax that uses simple, static, data object methods to perform simple data tasks with one line of code are common in scripting languages and are a good match for folks working in PHP/Python, etc. Seems like Microsoft has taken great advantage of .NET 4.0’s dynamic typing to provide this sort of interface for row iteration where each row has properties for each field. FWIW, all the examples demonstrate using local SQL Compact files - I was unable to get a SQL Server connection string to work with the Database class (the connection string wasn’t accepted). However, since the code in the page is still plain old .NET, you can easily use standard ADO.NET code or even LINQ or Entity Framework models that are created outside of WebMatrix in separate assemblies as required. The good the bad the obnoxious - It’s still .NET The beauty (or curse depending on how you look at it :)) of Razor and the compilation model is that, behind it all, it’s still .NET. Although the syntax may look foreign, it’s still all .NET behind the scenes. You can easily access existing tools, helpers, and utilities simply by adding them to the project as references or to the bin folder. Razor automatically recognizes any assembly reference from assemblies in the bin folder. In the default configuration, Microsoft provides a host of helper functions in a Microsoft.WebPages assembly (check it out in the ASP.NET temp folder for your application), which includes a host of HTML Helpers. If you’ve used ASP.NET MVC before, a lot of the helpers should look familiar. Documentation at the moment is sketchy-there’s a very rough API reference you can check out here: http://www.asp.net/webmatrix/tutorials/asp-net-web-pages-api-reference Who needs WebMatrix? Uhm… good Question Clearly Microsoft is trying hard to create an environment with WebMatrix that is easy to use for newbie developers. The goal seems to be simplicity in providing a minimal development environment and an easy-to-use script engine/language that makes it easy to get started with. There’s also some focus on community features that can be used as starting points, such as Web Gallery applications and templates. The community features in particular are very nice and something that would be nice to eventually see in Visual Studio as well. The question is whether this is too little too late. Developers who have been clamoring for a simpler development environment on the .NET stack have mostly left for other simpler platforms like PHP or Python which are catering to the down and dirty developer. Microsoft will be hard pressed to win those folks-and other hardcore PHP developers-back. Regardless of how much you dress up a script engine fronted by the .NET Framework, it’s still the .NET Framework and all the complexity that drives it. While .NET is a fine solution in its breadth and features once you get a basic handle on the core features, the bar of entry to being productive with the .NET Framework is still pretty high. The MVC style helpers Microsoft provides are a good step in the right direction, but I suspect it’s not enough to shield new developers from having to delve much deeper into the Framework to get even basic applications built. Razor and its helpers is trying to make .NET more accessible but the reality is that in order to do useful stuff that goes beyond the handful of simple helpers you still are going to have to write some C# or VB or other .NET code. If the target is a hobby/amateur/non-programmer the learning curve isn’t made any easier by WebMatrix it’s just been shifted a tad bit further along in your development endeavor when you run out of canned components that are supplied either by Microsoft or the community. The database helpers are interesting and actually I’ve heard a lot of discussion from various developers who’ve been resisting .NET for a really long time perking up at the prospect of easier data access in .NET than the ridiculous amount of code it takes to do even simple data access with raw ADO.NET. It seems sad that such a simple concept and implementation should trigger this sort of response (especially since it’s practically trivial to create helpers like these or pick them up from countless libraries available), but there it is. It also shows that there are plenty of developers out there who are more interested in ‘getting stuff done’ easily than necessarily following the latest and greatest practices which are overkill for many development scenarios. Sometimes it seems that all of .NET is focused on the big life changing issues of development, rather than the bread and butter scenarios that many developers are interested in to get their work accomplished. And that in the end may be WebMatrix’s main raison d'être: To bring some focus back at Microsoft that simpler and more high level solutions are actually needed to appeal to the non-high end developers as well as providing the necessary tools for the high end developers who want to follow the latest and greatest trends. The current version of WebMatrix hits many sweet spots, but it also feels like it has a long way to go before it really can be a tool that a beginning developer or an accomplished developer can feel comfortable with. Although there are some really good ideas in the environment (like the gallery for downloading apps and components) which would be a great addition for Visual Studio as well, the rest of the development environment just feels like crippleware with required functionality missing especially debugging and Intellisense, but also general editor support. It’s not clear whether these are because the product is still in an early alpha release or whether it’s simply designed that way to be a really limited development environment. While simple can be good, nobody wants to feel left out when it comes to necessary tool support and WebMatrix just has that left out feeling to it. If anything WebMatrix’s technology pieces (which are really independent of the WebMatrix product) are what are interesting to developers in general. The compact IIS implementation is a nice improvement for development scenarios and SQL Compact 4.0 seems to address a lot of concerns that people have had and have complained about for some time with previous SQL Compact implementations. By far the most interesting and useful technology though seems to be the Razor view engine for its light weight implementation and it’s decoupling from the ASP.NET/HTTP pipeline to provide a standalone scripting/view engine that is pluggable. The first winner of this is going to be ASP.NET MVC which can now have a cleaner view model that isn’t inconsistent due to the baggage of non-implemented WebForms features that don’t work in MVC. But I expect that Razor will end up in many other applications as a scripting and code generation engine eventually. Visual Studio integration for Razor is currently missing, but is promised for a later release. The ASP.NET MVC team has already mentioned that Razor will eventually become the default MVC view engine, which will guarantee continued growth and development of this tool along those lines. And the Razor engine and support tools actually inherit many of the features that MVC pioneered, so there’s some synergy flowing both ways between Razor and MVC. As an existing ASP.NET developer who’s already familiar with Visual Studio and ASP.NET development, the WebMatrix IDE doesn’t give you anything that you want. The tools provided are minimal and provide nothing that you can’t get in Visual Studio today, except the minimal Razor syntax highlighting, so there’s little need to take a step back. With Visual Studio integration coming later there’s little reason to look at WebMatrix for tooling. It’s good to see that Microsoft is giving some thought about the ease of use of .NET as a platform For so many years, we’ve been piling on more and more new features without trying to take a step back and see how complicated the development/configuration/deployment process has become. Sometimes it’s good to take a step - or several steps - back and take another look and realize just how far we’ve come. WebMatrix is one of those reminders and one that likely will result in some positive changes on the platform as a whole. © Rick Strahl, West Wind Technologies, 2005-2010Posted in ASP.NET   IIS7  

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• Could not resolve <fx:Script> to a component implementation.

  - by seref

  Hi, i created project with flexmojos maven archtype..i used flexmojos:flexbuilder and compile/run with FlashBuilder 4 everything is okay but when i try to compile project with flexmojos i got following error: [ERROR] Z:....\src\main\flex\Main.mxml:[6,-1] Could not resolve < fx:Script to a component implementation. [INFO] BUILD FAILURE my mxml: <?xml version="1.0" encoding="utf-8"?> <s:Application xmlns:fx="http://ns.adobe.com/mxml/2009" xmlns:s="library://ns.adobe.com/flex/spark" xmlns:mx="library://ns.adobe.com/flex/mx" width="100%" height="100%" creationComplete="application1_creationCompleteHandler(event)"> <fx:Script> <![CDATA[ import mx.controls.Alert; import mx.events.FlexEvent; protected function application1_creationCompleteHandler(event:FlexEvent):void { Alert.show("success!!!!") } ]]></fx:Script> </s:Application> pom.xml like: ...... <packaging>swf</packaging> ...... <properties> <flex-sdk.version>4.1.0.16076</flex-sdk.version> <flexmojos.version>3.8</flexmojos.version> </properties> ...... <build> <sourceDirectory>src/main/flex</sourceDirectory> <testSourceDirectory>src/test/flex</testSourceDirectory> <plugins> <plugin> <groupId>org.sonatype.flexmojos</groupId> <artifactId>flexmojos-maven-plugin</artifactId> <version>${flexmojos.version}</version> <extensions>true</extensions> <dependencies> <dependency> <groupId>com.adobe.flex</groupId> <artifactId>compiler</artifactId> <version>${flex-sdk.version}</version> <type>pom</type> </dependency> </dependencies> <configuration> <compiledLocales> <locale>en_US</locale> </compiledLocales> <mergeResourceBundle>true</mergeResourceBundle> <accessible>true</accessible> <optimize>true</optimize> <targetPlayer>10.0.0</targetPlayer> <showWarnings>true</showWarnings> <linkReport>true</linkReport> </configuration> </plugin> </plugins> </build> <dependencies> <!-- Flex framework resource bundles --> <dependency> <groupId>com.adobe.flex.framework</groupId> <artifactId>flex-framework</artifactId> <version>${flex-sdk.version}</version> <type>pom</type> </dependency> <!-- Include unit test dependencies. --> <dependency> <groupId>com.adobe.flexunit</groupId> <artifactId>flexunit</artifactId> <version>4.0-rc-1</version> <type>swc</type> <scope>test</scope> </dependency> </dependencies> ....... maven output compiler config : INFO] Flex compiler configurations: -compiler.external-library-path C:\...\.m2\repository\com\adobe\flex \framework\playerglobal\4.1.0.16076\10.0\playerglobal.swc -compiler.include-libraries= -compiler.library-path C:\...\.m2\repository\com\adobe\flex\framework \datavisualization\4.1.0.16076\datavisualization-4.1.0.16076.swc C:\... \.m2\repository\com\adobe\flex\framework\flash-integration \4.1.0.16076\flash-integration-4.1.0.16076.swc C:\...\.m2\repository \com\adobe\flex\framework\flex\4.1.0.16076\flex-4.1.0.16076.swc C:\... \.m2\repository\com\adobe\flex\framework\framework \4.1.0.16076\framework-4.1.0.16076.swc C:\...\.m2\repository\com\adobe \flex\framework\osmf\4.1.0.16076\osmf-4.1.0.16076.swc C:\... \.m2\repository\com\adobe\flex\framework\rpc \4.1.0.16076\rpc-4.1.0.16076.swc C:\...\.m2\repository\com\adobe\flex \framework\spark\4.1.0.16076\spark-4.1.0.16076.swc C:\... \.m2\repository\com\adobe\flex\framework\sparkskins \4.1.0.16076\sparkskins-4.1.0.16076.swc C:\...\.m2\repository\com\adobe \flex\framework\textLayout\4.1.0.16076\textLayout-4.1.0.16076.swc C: \...\.m2\repository\com\adobe\flex\framework\utilities \4.1.0.16076\utilities-4.1.0.16076.swc C:\...\.m2\repository\com\adobe \flex\framework\datavisualization \4.1.0.16076\datavisualization-4.1.0.16076-en_US.rb.swc C:\... \.m2\repository\com\adobe\flex\framework\framework \4.1.0.16076\framework-4.1.0.16076-en_US.rb.swc C:\...\.m2\repository \com\adobe\flex\framework\osmf\4.1.0.16076\osmf-4.1.0.16076- en_US.rb.swc C:\...\.m2\repository\com\adobe\flex\framework\rpc \4.1.0.16076\rpc-4.1.0.16076-en_US.rb.swc C:\...\.m2\repository\com \adobe\flex\framework\spark\4.1.0.16076\spark-4.1.0.16076-en_US.rb.swc C:\...\.m2\repository\com\adobe\flex\framework\textLayout \4.1.0.16076\textLayout-4.1.0.16076-en_US.rb.swc C:\...\.m2\repository \com\adobe\flex\framework\flash-integration\4.1.0.16076\flash- integration-4.1.0.16076-en_US.rb.swc C:\...\.m2\repository\com\adobe \flex\framework\playerglobal\4.1.0.16076\playerglobal-4.1.0.16076- en_US.rb.swc -compiler.theme Z:\.....\target\classes\configs\themes\Spark \spark.css -compiler.accessible=true -compiler.allow-source-path-overlap=false -compiler.as3=true -compiler.debug=false -compiler.es=false -compiler.fonts.managers flash.fonts.JREFontManager flash.fonts.BatikFontManager flash.fonts.AFEFontManager flash.fonts.CFFFontManager -compiler.fonts.local-fonts-snapshot Z:\.....\target\classes \fonts.ser -compiler.keep-generated-actionscript=false -licenses.license flashbuilder4 952309948800588759250406 -licenses.license flexbuilder4.displayedStartPageAtLeastOneTime true -compiler.locale en_US -compiler.optimize=true -compiler.source-path Z:\.....\src\main\flex -compiler.strict=true -use-network=true -compiler.verbose-stacktraces=false -compiler.actionscript-file-encoding UTF-8 -target-player 10.0.0 -default-background-color 8821927 -default-frame-rate 24 -default-script-limits 1000 60 -default-size 500 375 -compiler.headless-server=false -compiler.keep-all-type-selectors=false -compiler.use-resource-bundle-metadata=true -metadata.date Fri Mar 04 14:04:37 EET 2011 -metadata.localized-title Main x-default -verify-digests=true -compiler.namespaces.namespace+=http://ns.adobe.com/mxml/2009,Z:\..... \target\classes\config-4.1.0.16076\mxml-2009-manifest.xml -compiler.namespaces.namespace+=library://ns.adobe.com/flex/spark,Z: \.....\target\classes\config-4.1.0.16076\spark-manifest.xml -compiler.namespaces.namespace+=library://ns.adobe.com/flex/mx,Z:\..... \target\classes\config-4.1.0.16076\mx-manifest.xml -compiler.namespaces.namespace+=http://www.adobe.com/2006/mxml,Z:\..... \PozitronUI\target\classes\config-4.1.0.16076\mxml-manifest.xml - static-link-runtime-shared-libraries=false -load-config= -metadata.language+=en_US any help... regards,

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• Impossible to do POSTs with appengine-jruby/RoR: Reflection is not allowed

  - by Joel Cuevas

  I'm trying to build a site with RoR on Google App Engine. I'm using the google-appengine gem (http://appengine-jruby.googlecode.com) and following the instructions in (http://gist.github.com/268192). The problem is that I can't submit ANY form! I've already tried this in two diferent clean Win 7 Pro envs and the result is the same. After install Ruby 1.8.6 (One-Click Installer): 1. gem update --system 2. gem install rails 3. gem install google-appengine 4. gem install rails_dm_datastore 5. gem install activerecord-nulldb-adapter 6. curl -O http://appengine-jruby.googlecode.com/hg/demos/rails2/rails2_appengine.rb 7. ruby rails2_appengine.rb (previously downloaded) 8. rails myproj 9. chmod myproj 10. ruby script/generate dd_model MyModel f1:string f2:float f3:float f4:float f5:integer f6:integer f7:integer -f 11. ruby script/generate scaffold MyModel f1:string f2:float f3:float f4:float f5:integer f6:integer f7:integer -f --skip-migration 12. dev_appserver.rb -p 3000 . At this point, I manually test the scaffold in (http://localhost:3000/my_models). The index is OK, then I create a new registry with the generated form, everything's fine, but when I try to create a second one, I get a "java.lang.RuntimeException: DummyDynamicScope should never be used for backref storage" in the console. As far as I read this is a won't-fix behavior in JRuby 1.4.1, but it's converted to a debug only warning in 1.5.0, so I proceed to install the pre release. 13. gem install appengine-jruby-jars --pre With this, that exception is solved and everything works great... until I move the project to the GAE server. 14. ruby appcfg.rb update . And now, in (http://myproj.appspot.com/my_models), again, the index is fine, also the new form, but in the moment that I submit it with valid data, I get a 500 error: "java.lang.IllegalAccessException: Reflection is not allowed on public int". As I said, this behavior is not present in the local SDK. In both cases, I'm completely unable to post anything. This is what I have right now in the GAE environment: Ruby version 1.8.7 (java) RubyGems disabled Rack version 1.1 Rails version 2.3.5 Action Pack version 2.3.5 Active Support version 2.3.5 DataMapper version 0.10.2 Environment production JRuby Runtime version 1.5.0.pre JRuby-Rack version 0.9.7 AppEngine SDK version Google App Engine/1.3.3 AppEngine APIs version 0.0.15 And this are my intalled gems: actionmailer (2.3.5) actionpack (2.3.5) activerecord (2.3.5) activerecord-nulldb-adapter (0.2.0) activeresource (2.3.5) activesupport (2.3.5) addressable (2.1.2) appengine-apis (0.0.15) appengine-jruby-jars (0.0.8.pre, 0.0.7) appengine-rack (0.0.8) appengine-sdk (1.3.3.1) appengine-tools (0.0.12) bundler08 (0.8.5) dm-appengine (0.0.8) dm-ar-finders (0.10.2) dm-core (0.10.2) dm-timestamps (0.10.2) dm-validations (0.10.2) extlib (0.9.14) fxri (0.3.7, 0.3.6) google-appengine (0.0.12) hpricot (0.8.2 x86-mswin32, 0.6 mswin32) jruby-rack (0.9.8, 0.9.7) log4r (1.1.7, 1.0.5) rack (1.1.0, 1.0.1) rails (2.3.5) rails_appengine (0.0.3) rails_dm_datastore (0.2.9) rake (0.8.7, 0.7.3) rubygems-update (1.3.7, 1.3.6) rubyzip (0.9.4) sources (0.0.1) win32-api (1.4.6 x86-mswin32-60, 1.0.4 mswin32) win32-clipboard (0.5.2, 0.4.3) win32-dir (0.3.6, 0.3.2) win32-eventlog (0.5.2, 0.4.6) win32-file (0.6.3, 0.5.4) win32-file-stat (1.3.4, 1.2.7) win32-process (0.6.2, 0.5.3) win32-sapi (0.1.5, 0.1.4) win32-sound (0.4.2, 0.4.1) windows-api (0.4.0, 0.2.0) windows-pr (1.0.9, 0.7.2) I'm unable to attach the full logs of the exceptions because of the character limits, but I can provide them under request. Here's an abstract of them: DummyDynamicScope (dev and prod envs): 14-may-2010 7:18:40 com.google.appengine.tools.development.ApiProxyLocalImpl log SEVERE: [1273821520195000] javax.servlet.ServletContext log: Application Error java.lang.RuntimeException: DummyDynamicScope should never be used for backref storage at org.jruby.runtime.scope.DummyDynamicScope.getBackRef(DummyDynamicScope.java:49) at org.jruby.RubyRegexp.updateBackRef(RubyRegexp.java:1404) at org.jruby.RubyRegexp.updateBackRef(RubyRegexp.java:1396) at org.jruby.RubyRegexp.search(RubyRegexp.java:1386) at org.jruby.RubyRegexp.op_match(RubyRegexp.java:1301) at org.jruby.RubyString.op_match(RubyString.java:1446) at org.jruby.RubyString$i_method_1_0$RUBYINVOKER$op_match.call(org/jruby/RubyString$i_method_1_0$RUBYINVOKER$op_match.gen) at org.jruby.internal.runtime.methods.JavaMethod$JavaMethodOneOrN.call(JavaMethod.java:721) at org.jruby.RubyClass.finvoke(RubyClass.java:472) at org.jruby.RubyObject.send(RubyObject.java:1442) at org.jruby.RubyObject$i_method_multi$RUBYINVOKER$send.call(org/jruby/RubyObject$i_method_multi$RUBYINVOKER$send.gen) at org.jruby.internal.runtime.methods.JavaMethod$JavaMethodZeroOrOneOrTwoOrNBlock.call(JavaMethod.java:276) at org.jruby.runtime.callsite.CachingCallSite.cacheAndCall(CachingCallSite.java:330) at org.jruby.runtime.callsite.CachingCallSite.call(CachingCallSite.java:189) at ruby.jit.ruby.C_3a_.Desarrollo.AppEngine.gorgory.WEB_minus_INF.lib.gems_dot_jar.bundler_gems.jruby.$1_dot_8.gems.dm_minus_validations_minus_0_dot_10_dot_2.lib.dm_minus_validations.validators.numeric_validator.validate_with_comparison at ruby.jit.ruby.C_3a_.Desarrollo.AppEngine.gorgory.WEB_minus_INF.lib.gems_dot_jar.bundler_gems.jruby.$1_dot_8.gems.dm_minus_validations_minus_0_dot_10_dot_2.lib.dm_minus_validations.validators.numeric_validator.validate_with_comparison at org.jruby.internal.runtime.methods.JittedMethod.call(JittedMethod.java:102) at org.jruby.internal.runtime.methods.DefaultMethod.call(DefaultMethod.java:144) at org.jruby.runtime.callsite.CachingCallSite.cacheAndCall(CachingCallSite.java:280) at org.jruby.runtime.callsite.CachingCallSite.call(CachingCallSite.java:69) at org.jruby.ast.FCallManyArgsNode.interpret(FCallManyArgsNode.java:60) at org.jruby.ast.NewlineNode.interpret(NewlineNode.java:104) at org.jruby.internal.runtime.methods.InterpretedMethod.call(InterpretedMethod.java:229) at org.jruby.internal.runtime.methods.DefaultMethod.call(DefaultMethod.java:193) at org.jruby.RubyClass.finvoke(RubyClass.java:491) at org.jruby.RubyObject.send(RubyObject.java:1448) at org.jruby.RubyObject$i_method_multi$RUBYINVOKER$send.call(org/jruby/RubyObject$i_method_multi$RUBYINVOKER$send.gen) at org.jruby.internal.runtime.methods.JavaMethod$JavaMethodZeroOrOneOrTwoOrThreeOrNBlock.call(JavaMethod.java:293) at org.jruby.runtime.callsite.CachingCallSite.cacheAndCall(CachingCallSite.java:350) at org.jruby.runtime.callsite.CachingCallSite.call(CachingCallSite.java:229) at ruby.jit.ruby.C_3a_.Desarrollo.AppEngine.gorgory.WEB_minus_INF.lib.gems_dot_jar.bundler_gems.jruby.$1_dot_8.gems.dm_minus_validations_minus_0_dot_10_dot_2.lib.dm_minus_validations.validators.numeric_validator.validate_with28985350_50 at ruby.jit.ruby.C_3a_.Desarrollo.AppEngine.gorgory.WEB_minus_INF.lib.gems_dot_jar.bundler_gems.jruby.$1_dot_8.gems.dm_minus_validations_minus_0_dot_10_dot_2.lib.dm_minus_validations.validators.numeric_validator.validate_with28985350_50 at org.jruby.internal.runtime.methods.JittedMethod.call(JittedMethod.java:221) at org.jruby.internal.runtime.methods.DefaultMethod.call(DefaultMethod.java:201) at org.jruby.runtime.callsite.CachingCallSite.call(CachingCallSite.java:227) at org.jruby.ast.FCallThreeArgNode.interpret(FCallThreeArgNode.java:40) Reflection (only prod env): Java::JavaLang::SecurityException (java.lang.IllegalAccessException: Reflection is not allowed on public int java.lang.String$CaseInsensitiveComparator.compare(java.lang.String,java.lang.String)): com.google.appengine.runtime.Request.process-92563a0605f433ea(Request.java) java.lang.reflect.AccessibleObject.setAccessible(AccessibleObject.java:40) org.jruby.javasupport.JavaMethod.<init>(JavaMethod.java:176) org.jruby.javasupport.JavaMethod.create(JavaMethod.java:183) org.jruby.java.invokers.MethodInvoker.createCallable(MethodInvoker.java:23) org.jruby.java.invokers.RubyToJavaInvoker.<init>(RubyToJavaInvoker.java:63) org.jruby.java.invokers.MethodInvoker.<init>(MethodInvoker.java:13) org.jruby.java.invokers.InstanceMethodInvoker.<init>(InstanceMethodInvoker.java:15) org.jruby.javasupport.JavaClass$InstanceMethodInvokerInstaller.install(JavaClass.java:339) org.jruby.javasupport.JavaClass.installClassMethods(JavaClass.java:723) org.jruby.javasupport.JavaClass.setupProxy(JavaClass.java:586) org.jruby.javasupport.Java.createProxyClass(Java.java:506) org.jruby.javasupport.Java.getProxyClass(Java.java:445) org.jruby.javasupport.Java.getInstance(Java.java:354) org.jruby.javasupport.JavaUtil.convertJavaToUsableRubyObject(JavaUtil.java:143) org.jruby.javasupport.JavaClass$ConstantField.install(JavaClass.java:360) org.jruby.javasupport.JavaClass.installClassFields(JavaClass.java:711) org.jruby.javasupport.JavaClass.setupProxy(JavaClass.java:585) org.jruby.javasupport.Java.createProxyClass(Java.java:506) org.jruby.javasupport.Java.getProxyClass(Java.java:445) org.jruby.javasupport.Java.getProxyOrPackageUnderPackage(Java.java:885) org.jruby.javasupport.Java.get_proxy_or_package_under_package(Java.java:918) org.jruby.javasupport.JavaUtilities.get_proxy_or_package_under_package(JavaUtilities.java:54) org.jruby.javasupport.JavaUtilities$s_method_2_0$RUBYINVOKER$get_proxy_or_package_under_package.call(org/jruby/javasupport/JavaUtilities$s_method_2_0$RUBYINVOKER$get_proxy_or_package_under_package.gen:65535) org.jruby.runtime.callsite.CachingCallSite.cacheAndCall(CachingCallSite.java:329) org.jruby.runtime.callsite.CachingCallSite.call(CachingCallSite.java:188) org.jruby.ast.CallTwoArgNode.interpret(CallTwoArgNode.java:59) org.jruby.ast.NewlineNode.interpret(NewlineNode.java:104) org.jruby.ast.BlockNode.interpret(BlockNode.java:71) org.jruby.internal.runtime.methods.InterpretedMethod.call(InterpretedMethod.java:113) org.jruby.internal.runtime.methods.DefaultMethod.call(DefaultMethod.java:138) org.jruby.javasupport.util.RuntimeHelpers$MethodMissingMethod.call(RuntimeHelpers.java:389) org.jruby.internal.runtime.methods.DynamicMethod.call(DynamicMethod.java:182) What should I do now? Any hint would be wellcome. Thanks!

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• Async ignored on AJAX requests on Nginx server

  - by eComEvo

  Despite sending an async request to the server over AJAX, the server will not respond until the previous unrelated request has finished. The following code is only broken in this way on Nginx, but runs perfectly on Apache. This call will start a background process and it waits for it to complete so it can display the final result. $.ajax({ type: 'GET', async: true, url: $(this).data('route'), data: $('input[name=data]').val(), dataType: 'json', success: function (data) { /* do stuff */} error: function (data) { /* handle errors */} }); The below is called after the above, which on Apache requires 100ms to execute and repeats itself, showing progress for data being written in the background: checkStatusInterval = setInterval(function () { $.ajax({ type: 'GET', async: false, cache: false, url: '/process-status?process=' + currentElement.attr('id'), dataType: 'json', success: function (data) { /* update progress bar and status message */ } }); }, 1000); Unfortunately, when this script is run from nginx, the above progress request never even finishes a single request until the first AJAX request that sent the data is done. If I change the async to TRUE in the above, it executes one every interval, but none of them complete until that very first AJAX request finishes. Here is the main nginx conf file: #user nobody; worker_processes 1; #error_log logs/error.log; #error_log logs/error.log notice; #error_log logs/error.log info; #pid logs/nginx.pid; events { worker_connections 1024; } http { include mime.types; default_type application/octet-stream; server_names_hash_bucket_size 64; # configure temporary paths # nginx is started with param -p, setting nginx path to serverpack installdir fastcgi_temp_path temp/fastcgi; uwsgi_temp_path temp/uwsgi; scgi_temp_path temp/scgi; client_body_temp_path temp/client-body 1 2; proxy_temp_path temp/proxy; log_format main '$remote_addr - $remote_user [$time_local] "$request" ' '$status $body_bytes_sent "$http_referer" ' '"$http_user_agent" "$http_x_forwarded_for"'; #access_log logs/access.log main; # Sendfile copies data between one FD and other from within the kernel. # More efficient than read() + write(), since the requires transferring data to and from the user space. sendfile on; # Tcp_nopush causes nginx to attempt to send its HTTP response head in one packet, # instead of using partial frames. This is useful for prepending headers before calling sendfile, # or for throughput optimization. tcp_nopush on; # don't buffer data-sends (disable Nagle algorithm). Good for sending frequent small bursts of data in real time. tcp_nodelay on; types_hash_max_size 2048; # Timeout for keep-alive connections. Server will close connections after this time. keepalive_timeout 90; # Number of requests a client can make over the keep-alive connection. This is set high for testing. keepalive_requests 100000; # allow the server to close the connection after a client stops responding. Frees up socket-associated memory. reset_timedout_connection on; # send the client a "request timed out" if the body is not loaded by this time. Default 60. client_header_timeout 20; client_body_timeout 60; # If the client stops reading data, free up the stale client connection after this much time. Default 60. send_timeout 60; # Size Limits client_body_buffer_size 64k; client_header_buffer_size 4k; client_max_body_size 8M; # FastCGI fastcgi_connect_timeout 60; fastcgi_send_timeout 120; fastcgi_read_timeout 300; # default: 60 secs; when step debugging with XDEBUG, you need to increase this value fastcgi_buffer_size 64k; fastcgi_buffers 4 64k; fastcgi_busy_buffers_size 128k; fastcgi_temp_file_write_size 128k; # Caches information about open FDs, freqently accessed files. open_file_cache max=200000 inactive=20s; open_file_cache_valid 30s; open_file_cache_min_uses 2; open_file_cache_errors on; # Turn on gzip output compression to save bandwidth. # http://wiki.nginx.org/HttpGzipModule gzip on; gzip_disable "MSIE [1-6]\.(?!.*SV1)"; gzip_http_version 1.1; gzip_vary on; gzip_proxied any; #gzip_proxied expired no-cache no-store private auth; gzip_comp_level 6; gzip_buffers 16 8k; gzip_types text/plain text/css application/json application/x-javascript text/xml application/xml application/xml+rss text/javascript application/javascript; # show all files and folders autoindex on; server { # access from localhost only listen 127.0.0.1:80; server_name localhost; root www; # the following default "catch-all" configuration, allows access to the server from outside. # please ensure your firewall allows access to tcp/port 80. check your "skype" config. # listen 80; # server_name _; log_not_found off; charset utf-8; access_log logs/access.log main; # handle files in the root path /www location / { index index.php index.html index.htm; } #error_page 404 /404.html; # redirect server error pages to the static page /50x.html # error_page 500 502 503 504 /50x.html; location = /50x.html { root www; } # pass the PHP scripts to FastCGI server listening on 127.0.0.1:9100 # location ~ \.php$ { try_files $uri =404; fastcgi_pass 127.0.0.1:9100; fastcgi_index index.php; fastcgi_param SCRIPT_FILENAME $document_root$fastcgi_script_name; include fastcgi_params; } # add expire headers location ~* ^.+.(gif|ico|jpg|jpeg|png|flv|swf|pdf|mp3|mp4|xml|txt|js|css)$ { expires 30d; } # deny access to .htaccess files (if Apache's document root concurs with nginx's one) # deny access to git & svn repositories location ~ /(\.ht|\.git|\.svn) { deny all; } } # include config files of "enabled" domains include domains-enabled/*.conf; } Here is the enabled domain conf file: access_log off; access_log C:/server/www/test.dev/logs/access.log; error_log C:/server/www/test.dev/logs/error.log; # HTTP Server server { listen 127.0.0.1:80; server_name test.dev; root C:/server/www/test.dev/public; index index.php; rewrite_log on; default_type application/octet-stream; #include /etc/nginx/mime.types; # Include common configurations. include domains-common/location.conf; } # HTTPS server server { listen 443 ssl; server_name test.dev; root C:/server/www/test.dev/public; index index.php; rewrite_log on; default_type application/octet-stream; #include /etc/nginx/mime.types; # Include common configurations. include domains-common/location.conf; include domains-common/ssl.conf; } Contents of ssl.conf: # OpenSSL for HTTPS connections. ssl on; ssl_certificate C:/server/bin/openssl/certs/cert.pem; ssl_certificate_key C:/server/bin/openssl/certs/cert.key; ssl_session_timeout 5m; ssl_protocols SSLv3 TLSv1 TLSv1.1 TLSv1.2; ssl_ciphers HIGH:!aNULL:!MD5; ssl_prefer_server_ciphers on; # Pass the PHP scripts to FastCGI server listening on 127.0.0.1:9100 location ~ \.php$ { try_files $uri =404; fastcgi_param HTTPS on; fastcgi_pass 127.0.0.1:9100; fastcgi_index index.php; fastcgi_param SCRIPT_FILENAME $document_root$fastcgi_script_name; include fastcgi_params; } Contents of location.conf: # Remove trailing slash to please Laravel routing system. if (!-d $request_filename) { rewrite ^/(.+)/$ /$1 permanent; } location / { try_files $uri $uri/ /index.php?$query_string; } # We don't need .ht files with nginx. location ~ /(\.ht|\.git|\.svn) { deny all; } # Added cache headers for images. location ~* \.(png|jpg|jpeg|gif)$ { expires 30d; log_not_found off; } # Only 3 hours on CSS/JS to allow me to roll out fixes during early weeks. location ~* \.(js|css)$ { expires 3h; log_not_found off; } # Add expire headers. location ~* ^.+.(gif|ico|jpg|jpeg|png|flv|swf|pdf|mp3|mp4|xml|txt)$ { expires 30d; } # Pass the PHP scripts to FastCGI server listening on 127.0.0.1:9100 location ~ \.php$ { try_files $uri /index.php =404; fastcgi_index index.php; fastcgi_param SCRIPT_FILENAME $document_root$fastcgi_script_name; include fastcgi_params; fastcgi_pass 127.0.0.1:9100; } Any ideas where this is going wrong?

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• High Load mysql on Debian server stops every day. Why?

  - by Oleg Abrazhaev

  I have Debian server with 32 gb memory. And there is apache2, memcached and nginx on this server. Memory load always on maximum. Only 500m free. Most memory leak do MySql. Apache only 70 clients configured, other services small memory usage. When mysql use all memory it stops. And nothing works, need mysql reboot. Mysql configured use maximum 24 gb memory. I have hight weight InnoDB bases. (400000 rows, 30 gb). And on server multithread daemon, that makes many inserts in this tables, thats why InnoDB. There is my mysql config. [mysqld] # # * Basic Settings # default-time-zone = "+04:00" user = mysql pid-file = /var/run/mysqld/mysqld.pid socket = /var/run/mysqld/mysqld.sock port = 3306 basedir = /usr datadir = /var/lib/mysql tmpdir = /tmp language = /usr/share/mysql/english skip-external-locking default-time-zone='Europe/Moscow' # # Instead of skip-networking the default is now to listen only on # localhost which is more compatible and is not less secure. # # * Fine Tuning # #low_priority_updates = 1 concurrent_insert = ALWAYS wait_timeout = 600 interactive_timeout = 600 #normal key_buffer_size = 2024M #key_buffer_size = 1512M #70% hot cache key_cache_division_limit= 70 #16-32 max_allowed_packet = 32M #1-16M thread_stack = 8M #40-50 thread_cache_size = 50 #orderby groupby sort sort_buffer_size = 64M #same myisam_sort_buffer_size = 400M #temp table creates when group_by tmp_table_size = 3000M #tables in memory max_heap_table_size = 3000M #on disk open_files_limit = 10000 table_cache = 10000 join_buffer_size = 5M # This replaces the startup script and checks MyISAM tables if needed # the first time they are touched myisam-recover = BACKUP #myisam_use_mmap = 1 max_connections = 200 thread_concurrency = 8 # # * Query Cache Configuration # #more ignored query_cache_limit = 50M query_cache_size = 210M #on query cache query_cache_type = 1 # # * Logging and Replication # # Both location gets rotated by the cronjob. # Be aware that this log type is a performance killer. #log = /var/log/mysql/mysql.log # # Error logging goes to syslog. This is a Debian improvement :) # # Here you can see queries with especially long duration log_slow_queries = /var/log/mysql/mysql-slow.log long_query_time = 1 log-queries-not-using-indexes # # The following can be used as easy to replay backup logs or for replication. # note: if you are setting up a replication slave, see README.Debian about # other settings you may need to change. #server-id = 1 #log_bin = /var/log/mysql/mysql-bin.log server-id = 1 log-bin = /var/lib/mysql/mysql-bin #replicate-do-db = gate log-bin-index = /var/lib/mysql/mysql-bin.index log-error = /var/lib/mysql/mysql-bin.err relay-log = /var/lib/mysql/relay-bin relay-log-info-file = /var/lib/mysql/relay-bin.info relay-log-index = /var/lib/mysql/relay-bin.index binlog_do_db = 24avia expire_logs_days = 10 max_binlog_size = 100M read_buffer_size = 4024288 innodb_buffer_pool_size = 5000M innodb_flush_log_at_trx_commit = 2 innodb_thread_concurrency = 8 table_definition_cache = 2000 group_concat_max_len = 16M #binlog_do_db = gate #binlog_ignore_db = include_database_name # # * BerkeleyDB # # Using BerkeleyDB is now discouraged as its support will cease in 5.1.12. #skip-bdb # # * InnoDB # # InnoDB is enabled by default with a 10MB datafile in /var/lib/mysql/. # Read the manual for more InnoDB related options. There are many! # You might want to disable InnoDB to shrink the mysqld process by circa 100MB. #skip-innodb # # * Security Features # # Read the manual, too, if you want chroot! # chroot = /var/lib/mysql/ # # For generating SSL certificates I recommend the OpenSSL GUI "tinyca". # # ssl-ca=/etc/mysql/cacert.pem # ssl-cert=/etc/mysql/server-cert.pem # ssl-key=/etc/mysql/server-key.pem [mysqldump] quick quote-names max_allowed_packet = 500M [mysql] #no-auto-rehash # faster start of mysql but no tab completition [isamchk] key_buffer = 32M key_buffer_size = 512M # # * NDB Cluster # # See /usr/share/doc/mysql-server-*/README.Debian for more information. # # The following configuration is read by the NDB Data Nodes (ndbd processes) # not from the NDB Management Nodes (ndb_mgmd processes). # # [MYSQL_CLUSTER] # ndb-connectstring=127.0.0.1 # # * IMPORTANT: Additional settings that can override those from this file! # The files must end with '.cnf', otherwise they'll be ignored. # !includedir /etc/mysql/conf.d/ Please, help me make it stable. Memory used /etc/mysql # free total used free shared buffers cached Mem: 32930800 32766424 164376 0 139208 23829196 -/+ buffers/cache: 8798020 24132780 Swap: 33553328 44660 33508668 Maybe my problem not in memory, but MySQL stops every day. As you can see, cache memory free 24 gb. Thank to Michael Hampton? for correction. Load overage on server 3.5. Maybe hdd or another problem? Maybe my config not optimal for 30gb InnoDB ? I'm already try mysqltuner and tunung-primer.sh , but they marked all green. Mysqltuner output mysqltuner >> MySQLTuner 1.0.1 - Major Hayden <[email protected]> >> Bug reports, feature requests, and downloads at http://mysqltuner.com/ >> Run with '--help' for additional options and output filtering -------- General Statistics -------------------------------------------------- [--] Skipped version check for MySQLTuner script [OK] Currently running supported MySQL version 5.5.24-9-log [OK] Operating on 64-bit architecture -------- Storage Engine Statistics ------------------------------------------- [--] Status: -Archive -BDB -Federated +InnoDB -ISAM -NDBCluster [--] Data in MyISAM tables: 112G (Tables: 1528) [--] Data in InnoDB tables: 39G (Tables: 340) [--] Data in PERFORMANCE_SCHEMA tables: 0B (Tables: 17) [!!] Total fragmented tables: 344 -------- Performance Metrics ------------------------------------------------- [--] Up for: 8h 18m 33s (14M q [478.333 qps], 259K conn, TX: 9B, RX: 5B) [--] Reads / Writes: 84% / 16% [--] Total buffers: 10.5G global + 81.1M per thread (200 max threads) [OK] Maximum possible memory usage: 26.3G (83% of installed RAM) [OK] Slow queries: 1% (259K/14M) [!!] Highest connection usage: 100% (201/200) [OK] Key buffer size / total MyISAM indexes: 1.5G/5.6G [OK] Key buffer hit rate: 100.0% (6B cached / 1M reads) [OK] Query cache efficiency: 74.3% (8M cached / 11M selects) [OK] Query cache prunes per day: 0 [OK] Sorts requiring temporary tables: 0% (0 temp sorts / 247K sorts) [!!] Joins performed without indexes: 106025 [!!] Temporary tables created on disk: 49% (351K on disk / 715K total) [OK] Thread cache hit rate: 99% (249 created / 259K connections) [!!] Table cache hit rate: 15% (2K open / 13K opened) [OK] Open file limit used: 15% (3K/20K) [OK] Table locks acquired immediately: 99% (4M immediate / 4M locks) [!!] InnoDB data size / buffer pool: 39.4G/5.9G -------- Recommendations ----------------------------------------------------- General recommendations: Run OPTIMIZE TABLE to defragment tables for better performance MySQL started within last 24 hours - recommendations may be inaccurate Reduce or eliminate persistent connections to reduce connection usage Adjust your join queries to always utilize indexes Temporary table size is already large - reduce result set size Reduce your SELECT DISTINCT queries without LIMIT clauses Increase table_cache gradually to avoid file descriptor limits Variables to adjust: max_connections (> 200) wait_timeout (< 600) interactive_timeout (< 600) join_buffer_size (> 5.0M, or always use indexes with joins) table_cache (> 10000) innodb_buffer_pool_size (>= 39G) Mysql primer output -- MYSQL PERFORMANCE TUNING PRIMER -- - By: Matthew Montgomery - MySQL Version 5.5.24-9-log x86_64 Uptime = 0 days 8 hrs 20 min 50 sec Avg. qps = 478 Total Questions = 14369568 Threads Connected = 16 Warning: Server has not been running for at least 48hrs. It may not be safe to use these recommendations To find out more information on how each of these runtime variables effects performance visit: http://dev.mysql.com/doc/refman/5.5/en/server-system-variables.html Visit http://www.mysql.com/products/enterprise/advisors.html for info about MySQL's Enterprise Monitoring and Advisory Service SLOW QUERIES The slow query log is enabled. Current long_query_time = 1.000000 sec. You have 260626 out of 14369701 that take longer than 1.000000 sec. to complete Your long_query_time seems to be fine BINARY UPDATE LOG The binary update log is enabled Binlog sync is not enabled, you could loose binlog records during a server crash WORKER THREADS Current thread_cache_size = 50 Current threads_cached = 45 Current threads_per_sec = 0 Historic threads_per_sec = 0 Your thread_cache_size is fine MAX CONNECTIONS Current max_connections = 200 Current threads_connected = 11 Historic max_used_connections = 201 The number of used connections is 100% of the configured maximum. You should raise max_connections INNODB STATUS Current InnoDB index space = 214 M Current InnoDB data space = 39.40 G Current InnoDB buffer pool free = 0 % Current innodb_buffer_pool_size = 5.85 G Depending on how much space your innodb indexes take up it may be safe to increase this value to up to 2 / 3 of total system memory MEMORY USAGE Max Memory Ever Allocated : 23.46 G Configured Max Per-thread Buffers : 15.84 G Configured Max Global Buffers : 7.54 G Configured Max Memory Limit : 23.39 G Physical Memory : 31.40 G Max memory limit seem to be within acceptable norms KEY BUFFER Current MyISAM index space = 5.61 G Current key_buffer_size = 1.47 G Key cache miss rate is 1 : 5578 Key buffer free ratio = 77 % Your key_buffer_size seems to be fine QUERY CACHE Query cache is enabled Current query_cache_size = 200 M Current query_cache_used = 101 M Current query_cache_limit = 50 M Current Query cache Memory fill ratio = 50.59 % Current query_cache_min_res_unit = 4 K MySQL won't cache query results that are larger than query_cache_limit in size SORT OPERATIONS Current sort_buffer_size = 64 M Current read_rnd_buffer_size = 256 K Sort buffer seems to be fine JOINS Current join_buffer_size = 5.00 M You have had 106606 queries where a join could not use an index properly You have had 8 joins without keys that check for key usage after each row join_buffer_size >= 4 M This is not advised You should enable "log-queries-not-using-indexes" Then look for non indexed joins in the slow query log. OPEN FILES LIMIT Current open_files_limit = 20210 files The open_files_limit should typically be set to at least 2x-3x that of table_cache if you have heavy MyISAM usage. Your open_files_limit value seems to be fine TABLE CACHE Current table_open_cache = 10000 tables Current table_definition_cache = 2000 tables You have a total of 1910 tables You have 2151 open tables. The table_cache value seems to be fine TEMP TABLES Current max_heap_table_size = 2.92 G Current tmp_table_size = 2.92 G Of 366426 temp tables, 49% were created on disk Perhaps you should increase your tmp_table_size and/or max_heap_table_size to reduce the number of disk-based temporary tables Note! BLOB and TEXT columns are not allow in memory tables. If you are using these columns raising these values might not impact your ratio of on disk temp tables. TABLE SCANS Current read_buffer_size = 3 M Current table scan ratio = 2846 : 1 read_buffer_size seems to be fine TABLE LOCKING Current Lock Wait ratio = 1 : 185 You may benefit from selective use of InnoDB. If you have long running SELECT's against MyISAM tables and perform frequent updates consider setting 'low_priority_updates=1'

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• Char error C langauge

  - by Nadeem tabbaa

  i have a project for a course, i did almost everything but i have this error i dont know who to solve it... the project about doing our own shell some of them we have to write our code, others we will use the fork method.. this is the code, #include <sys/wait.h> #include <dirent.h> #include <limits.h> #include <errno.h> #include <stdlib.h> #include <string.h> #include<stdio.h> #include<fcntl.h> #include<unistd.h> #include<sys/stat.h> #include<sys/types.h> int main(int argc, char **argv) { pid_t pid; char str[21], *arg[10]; int x,status,number; system("clear"); while(1) { printf("Rshell>" ); fgets(str,21,stdin); x = 0; arg[x] = strtok(str, " \n\t"); while(arg[x]) arg[++x] = strtok(NULL, " \n\t"); if(NULL!=arg[0]) { if(strcasecmp(arg[0],"cat")==0) //done { int f=0,n; char l[1]; struct stat s; if(x!=2) { printf("Mismatch argument\n"); } /*if(access(arg[1],F_OK)) { printf("File Exist"); exit(1); } if(stat(arg[1],&s)<0) { printf("Stat ERROR"); exit(1); } if(S_ISREG(s.st_mode)<0) { printf("Not a Regular FILE"); exit(1); } if(geteuid()==s.st_uid) if(s.st_mode & S_IRUSR) f=1; else if(getegid()==s.st_gid) if(s.st_mode & S_IRGRP) f=1; else if(s.st_mode & S_IROTH) f=1; if(!f) { printf("Permission denied"); exit(1); }*/ f=open(arg[1],O_RDONLY); while((n=read(f,l,1))>0) write(1,l,n); } else if(strcasecmp(arg[0],"rm")==0) //done { if( unlink( arg[1] ) != 0 ) perror( "Error deleting file" ); else puts( "File successfully deleted" ); } else if(strcasecmp(arg[0],"rmdir")==0) //done { if( remove( arg[1] ) != 0 ) perror( "Error deleting Directory" ); else puts( "Directory successfully deleted" ); } else if(strcasecmp(arg[0],"ls")==0) //done { DIR *dir; struct dirent *dirent; char *where = NULL; //printf("x== %i\n",x); //printf("x== %s\n",arg[1]); //printf("x== %i\n",get_current_dir_name()); if (x == 1) where = get_current_dir_name(); else where = arg[1]; if (NULL == (dir = opendir(where))) { fprintf(stderr,"%d (%s) opendir %s failed\n", errno, strerror(errno), where); return 2; } while (NULL != (dirent = readdir(dir))) { printf("%s\n", dirent->d_name); } closedir(dir); } else if(strcasecmp(arg[0],"cp")==0) //not yet for Raed { FILE *from, *to; char ch; if(argc!=3) { printf("Usage: copy <source> <destination>\n"); exit(1); } /* open source file */ if((from = fopen(argv[1], "rb"))==NULL) { printf("Cannot open source file.\n"); exit(1); } /* open destination file */ if((to = fopen(argv[2], "wb"))==NULL) { printf("Cannot open destination file.\n"); exit(1); } /* copy the file */ while(!feof(from)) { ch = fgetc(from); if(ferror(from)) { printf("Error reading source file.\n"); exit(1); } if(!feof(from)) fputc(ch, to); if(ferror(to)) { printf("Error writing destination file.\n"); exit(1); } } if(fclose(from)==EOF) { printf("Error closing source file.\n"); exit(1); } if(fclose(to)==EOF) { printf("Error closing destination file.\n"); exit(1); } } else if(strcasecmp(arg[0],"mv")==0)//done { if( rename(arg[1],arg[2]) != 0 ) perror( "Error moving file" ); else puts( "File successfully moved" ); } else if(strcasecmp(arg[0],"hi")==0)//done { printf("hello\n"); } else if(strcasecmp(arg[0],"exit")==0) // done { return 0; } else if(strcasecmp(arg[0],"sleep")==0) // done { if(x==1) printf("plz enter the # seconds to sleep\n"); else sleep(atoi(arg[1])); } else if(strcmp(arg[0],"history")==0) // not done { FILE *infile; //char fname[40]; char line[100]; int lcount; ///* Read in the filename */ //printf("Enter the name of a ascii file: "); //fgets(History.txt, sizeof(fname), stdin); /* Open the file. If NULL is returned there was an error */ if((infile = fopen("History.txt", "r")) == NULL) { printf("Error Opening File.\n"); exit(1); } while( fgets(line, sizeof(line), infile) != NULL ) { /* Get each line from the infile */ lcount++; /* print the line number and data */ printf("Line %d: %s", lcount, line); } fclose(infile); /* Close the file */ writeHistory(arg); //write to txt file every new executed command //read from the file once the history command been called //if a command called not for the first time then just replace it to the end of the file } else if(strncmp(arg[0],"@",1)==0) // not done { //scripting files // read from the file command by command and executing them } else if(strcmp(arg[0],"type")==0) //not done { //if(x==1) //printf("plz enter the argument\n"); //else //type((arg[1])); } else { pid = fork( ); if (pid == 0) { execlp(arg[0], arg[0], arg[1], arg[2], NULL); printf ("EXEC Failed\n"); } else { wait(&status); if(strcmp(arg[0],"clear")!=0) { printf("status %04X\n",status); if(WIFEXITED(status)) printf("Normal termination, exit code %d\n", WEXITSTATUS(status)); else printf("Abnormal termination\n"); } } } } } } void writeHistory(char *arg[]) { FILE *file; file = fopen("History.txt","a+"); /* apend file (add text to a file or create a file if it does not exist.*/ int i =0; while(strcasecmp(arg[0],NULL)==0) { fprintf(file,"%s ",arg[i]); /*writes*/ } fprintf(file,"\n"); /*new line*/ fclose(file); /*done!*/ getchar(); /* pause and wait for key */ //return 0; } the thing is when i compile the code, this what it gives me /home/ugics/st255375/ICS431Labs/Project/Rshell.c: At top level: /home/ugics/st255375/ICS431Labs/Project/Rshell.c:264: warning: conflicting types for ‘writeHistory’ /home/ugics/st255375/ICS431Labs/Project/Rshell.c:217: note: previous implicit declaration of ‘writeHistory’ was here can any one help me??? thanks

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• Authoritative sources about Database vs. Flatfile decision

  - by FastAl

  <tldr>looking for a reference to a book or other undeniably authoritative source that gives reasons when you should choose a database vs. when you should choose other storage methods. I have provided an un-authoritative list of reasons about 2/3 of the way down this post.</tldr> I have a situation at my company where a database is being used where it would be better to use another solution (in this case, an auto-generated piece of source code that contains a static lookup table, searched by binary sort). Normally, a database would be an OK solution even though the problem does not require a database, e.g, none of the elements of ACID are needed, as it is read-only data, updated about every 3-5 years (also requiring other sourcecode changes), and fits in memory, and can be keyed into via binary search (a tad faster than db, but speed is not an issue). The problem is that this code runs on our enterprise server, but is shared with several PC platforms (some disconnected, some use a central DB, etc.), and parts of it are managed by multiple programming units, parts by the DBAs, parts even by mathematicians in another department, etc. These hit their own platform’s version of their databases (containing their own copy of the static data). What happens is that every implementation, every little change, something different goes wrong. There are many other issues as well. I can’t even use a flatfile, because one mode of running on our enterprise server does not have permission to read files (only databases, and of course, its own literal storage, e.g., in-source table). Of course, other parts of the system use databases in proper, less obscure manners; there is no problem with those parts. So why don’t we just change it? I don’t have administrative ability to force a change. But I’m affected because sometimes I have to help fix the problems, but mostly because it causes outages and tons of extra IT time by other programmers and d*mmit that makes me mad! The reason neither management, nor the designers of the system, can see the problem is that they propose a solution that won’t work: increase communication; implement more safeguards and standards; etc. But every time, in a different part of the already-pared-down but still multi-step processes, a few different diligent, hard-working, top performing IT personnel make a unique subtle error that causes it to fail, sometimes after the last round of testing! And in general these are not single-person failures, but understandable miscommunications. And communication at our company is actually better than most. People just don't think that's the case because they haven't dug into the matter. However, I have it on very good word from somebody with extensive formal study of sociology and psychology that the relatively small amount of less-than-proper database usage in this gigantic cross-platform multi-source, multi-language project is bureaucratically un-maintainable. Impossible. No chance. At least with Human Beings in the loop, and it can’t be automated. In addition, the management and developers who could change this, though intelligent and capable, don’t understand the rigidity of this ‘how humans are’ issue, and are not convincible on the matter. The reason putting the static data in sourcecode will solve the problem is, although the solution is less sexy than a database, it would function with no technical drawbacks; and since the sharing of sourcecode already works very well, you basically erase any database-related effort from this section of the project, along with all the drawbacks of it that are causing problems. OK, that’s the background, for the curious. I won’t be able to convince management that this is an unfixable sociological problem, and that the real solution is coding around these limits of human nature, just as you would code around a bug in a 3rd party component that you can’t change. So what I have to do is exploit the unsuitableness of the database solution, and not do it using logic, but rather authority. I am aware of many reasons, and posts on this site giving reasons for one over the other; I’m not looking for lists of reasons like these (although you can add a comment if I've miss a doozy): WHY USE A DATABASE? instead of flatfile/other DB vs. file: if you need... Random Read / Transparent search optimization Advanced / varied / customizable Searching and sorting capabilities Transaction/rollback Locks, semaphores Concurrency control / Shared users Security 1-many/m-m is easier Easy modification Scalability Load Balancing Random updates / inserts / deletes Advanced query Administrative control of design, etc. SQL / learning curve Debugging / Logging Centralized / Live Backup capabilities Cached queries / dvlp & cache execution plans Interleaved update/read Referential integrity, avoid redundant/missing/corrupt/out-of-sync data Reporting (from on olap or oltp db) / turnkey generation tools [Disadvantages:] Important to get right the first time - professional design - but only b/c it's meant to last s/w & h/w cost Usu. over a network, speed issue (best vs. best design vs. local=even then a separate process req's marshalling/netwk layers/inter-p comm) indicies and query processing can stand in the way of simple processing (vs. flatfile) WHY USE FLATFILE: If you only need... Sequential Row processing only Limited usage append only (no reading, no master key/update) Only Update the record you're reading (fixed length recs only) Too big to fit into memory If Local disk / read-ahead network connection Portability / small system Email / cut & Paste / store as document by novice - simple format Low design learning curve but high cost later WHY USE IN-MEMORY/TABLE (tables, arrays, etc.): if you need... Processing a single db/ff record that was imported Known size of data Static data if hardcoding the table Narrow, unchanging use (e.g., one program or proc) -includes a class that will be shared, but encapsulates its data manipulation Extreme speed needed / high transaction frequency Random access - but search is dependent on implementation Following are some other posts about the topic: http://stackoverflow.com/questions/1499239/database-vs-flat-text-file-what-are-some-technical-reasons-for-choosing-one-over http://stackoverflow.com/questions/332825/are-flat-file-databases-any-good http://stackoverflow.com/questions/2356851/database-vs-flat-files http://stackoverflow.com/questions/514455/databases-vs-plain-text/514530 What I’d like to know is if anybody could recommend a hard, authoritative source containing these reasons. I’m looking for a paper book I can buy, or a reputable website with whitepapers about the issue (e.g., Microsoft, IBM), not counting the user-generated content on those sites. This will have a greater change to elicit a change that I’m looking for: less wasted programmer time, and more reliable programs. Thanks very much for your help. You win a prize for reading such a large post!

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