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Introduction One hundred years ago, a recent college graduate created the US aluminum industry. He found an economic way to crack the primeval bond that makes the metal so special. Aluminum doesn't "rust" - oxidize progressively. Rather, when exposed to air, it instantly forms a gem-hard but microscopic oxide coating. This seals off oxygen from the metal underneath. Scratch it and the seal reforms instantly. That tight chemical bond is the reason the metal is not found in nature; it exists only as a compound. The bond had to be broken to free the shiny metal. But once it is free, that self-sealing protection, alongside a remarkable strength-to-weight ratio, offers a marvelous material. Until 100 years ago, finding a way to do that economically was a challenge that defied science. Then it was solved, in a remarkable coincidence, on both sides of the Atlantic. And the way was paved for a material that would become the second most widely used metal in the world. Until that time, aluminum was a rare and precious metal. But then two young scientists independently invented the process that took aluminum outside the laboratory and into applications ranging from pots and pans to spacecraft. Discovery of aluminum Other major metals - copper, iron, lead, gold, and silver - had been used for thousands of years. But aluminum was entirely unknown until 1782, when French chemist Antoine Lavoisier suspected its presence. It was not actually discovered until Sir Humphry Davy, in England, detected it in clay in 1807. It was first isolated as a free metal only in 1825, by Denmark's Hans Christian Oersted. For years, laboratory methods yielded aluminum only in tiny amounts costing about $1200/kg ($545 per lb) - $13,333/kg ($6000 per lb) in today's dollars. Methods developed by French scientist Henri Sainte-Claire Deville in 1854 cut the cost to about $200/kg ($90 per lb). King Frederick VII of Denmark had a helmet of aluminum and gold. And Emperor Napoleon III of France reportedly served his most exalted guests at a state dinner with aluminum plates. He left the lower ranks to dine from pure gold. Deville and others continued to reduce production costs. But the price of aluminum still rivaled silver - $31/kg ($14 per lb) versus about $35/kg ($16 per lb) for silver. Then in 1884, a 2.8-kg (6.25-1b), 230-mm-tall (9-in.-tall) aluminum pyramid was cast as the tip of the Washington Monument. It was probably the first architectural use of aluminum. And it is still there after more than a century. The attractive properties of aluminum were obvious: it is lightweight but strong, conducts heat and electricity, reflects light and other electromagnetic radiation, and resists corrosion. But its commercial promise was blocked by its cost. In 1885, the price of aluminum was $17.70/kg ($8 per lb). It stayed there for the next couple of years until the Hall-Heroult process brought it down sharply. Hall-Heroult process lowers aluminum costs The characteristic that kept aluminum hidden so long continued to thwart chemists. Aluminum clings to oxygen so tightly that it is never found as a free metal in nature, and it is hard to break that bond chemically. The best chemical methods of the 1880s required sodium, and sodium itself was expensive. At Oberlin College in Ohio, chemistry professor Frank Fanning Jewett advised his students that anyone who could make aluminum inexpensive would do humanity a favor and get rich as well.

Jan 3, 1987

By R. K. Mehta, S. K. Dhar, R. R. Mallepali

Introduction The incentives for developing a distributed form of the comminution model include: • for describing the heterogeneous comminution behavior of multiphase mineral systems (i.e. realistic ore bodies, such as complex sulfide ores consisting of chalcopyrite, galena, sphalerite, pyrite and siliceous minerals) and •for generating, for various locked and free particles, in situ, microscopic, breakage behavior such as the mode of fracture (intergranular vs. transgranular), the relative rates of breakage and the breakage distribution. However, the computational difficulty and the effort made in estimating a large number of parameters (Andrews and Mika, 1975; Wiegel, 1976; Herbst et al., 1985, 1988) do not justify the utility of this approach. Also, obtaining the results experimentally is very tedious and time consuming if using this approach. Although recent work in this area (Choi, Adel and Yoon, 1987; Mehta, Adel and Yoon, 1989, 1990) showed good results for liberated particles, the quality of the estimation and prediction was poor for locked particles. Attributing this to the complex breakage behavior of locked particles, it was concluded that additional constraints, and possibly modification in the breakage distribution functional form for locked particles, were necessary. However, further mathematical analysis revealed that the method previously used was in error. The purpose of this technical note is to present the correct solution methodology used in solving the general model equations of a multicomponent mineral system containing multiple classes of locked particles. Previous work and connection The general form of the comprehensive, distributed-component comminution model has been described as (Herbst et al., 1985) [d (m (0) -[ I - B ] S_ m(t) = A m(t) (1)] where [A] is the fundamental matrix/state companion matrix, commonly referred to, in the comminution literature, as the "mill matrix." Note that for a distributed-component comminution model, this is a full, general, real matrix, unlike a lower triangular matrix for a simple grinding model or a lumped-binary grinding model with one locked-particle class (Herbst et al., 1985, 1988). Also note that all of the underlined letters refer to matrices. Herbst et al. addressed the difficulty in arriving at a closed form of the analytical solution of Eq. (1) via similarity transformation [IT J inv (T)]]. This difficulty occurs for several reasons: •[A], as a general real matrix, may have complex eigenvalues and complex eigenvectors. •More importantly, there is no guarantee that a set of eigenvectors of [A] will turn out to be independent. An attempt was made to solve this problem by implementing a decomposition method. With this technique, the mill matrix was decomposed into a sum of lower and upper triangular matrices [(i.e., A = L + U)]. Then the solution was obtained using similarity transformation (Mehta, Adel and Yoon, 1989, 1990). [e(At) = e(L+U)t = e(Lt).e(Ut) _ [ TL Jt- (t) inv(T_L)] [-Tu Ju (t) inv(Tu)]] [But, unfortunately, e 1+U = e t . eu is true only for scalars and not for matrices (i.e., eL+u# eL- eU) unless they commute. In other words. eAt = e(L+u)t= ekt - eU' if LU = UL. Hence, this decomposition scheme must be replaced by a mathematically acceptable method known as the "splitting method"]

Jan 1, 1994

By L. P. Hendriks

Consolidated Murchison, the only antimony producer in South Africa, is situated in the northeastern part of Transvaal. Phalaborwa, an important copper/phosphate mining town is situated about 60 km to the east of Consolidated Murchison, on the border of the Kruger National Park. The mine is jointly controlled by Johannesburg Consolidated Investment Co. Ltd. (JCI) and Anglovaal Limited.

Jan 1, 1998

By Peter Zhang, Daniel Su, Bo Kim, Berk Tulu

This paper summarizes significant findings from the ongoing NIOSH Gas Well Stability research over the past decade; in particular, the important ground control engineering considerations. Longwall-induced surface and subsurface deformations may induce gas well casing deformations and stresses, depending on a few parameters such as overburden depth, overburden geology, topographic relief, strata dip, gas well setback distance, and a gas well casing cementing alternative. Overburden depth to the mining seam is one of the most influential factors; shallow overburden depth induces large horizontal displacement and resulting casing deformation and stress, while deep overburden depth induces large longwall-induced vertical pressure. Overburden geology is another important factor affecting the longwall- induced casing deformations and stresses; the large contrast of bending stiffness at the soft-to-hard rock interfaces tends to induce large casing deformations and stresses. Large surface topographic relief and the presence of soft-to-hard rock interfaces below the stream valley bottoms may shift longwall-induced subsurface gas well casing deformations closer to the surface, thus having less impact on underground. mining operations. Strata dip magnifies longwall-induced deformations by 20% per degree of strata dip for down dip longwall excavations. The gas well setback distance is also a critical geometric parameter; the effect of longwall-induced deformation decreases exponentially as the setback distance increases. Longwall-induced casing deformation and stress will be mitigated or totally uncoupled for uncemented production casing.

Jun 25, 2024

By Richard R. Klimpel

INTRODUCTION Previous papers by this author have discussed some of the difficulties involved with optimizing the performance of industrial flotation circuits, e.g. Klimpel (1988, 1989). The two major problems seem to be: 1) the inherent "systems" nature of the froth flotation process which leads to a great deal of self-compensation among the various operating options available and 2) the inherent variations in mineralogy, water chemistry, etc. that always seem to be present in industrial operations. In particular, the influence of reagent dosage has been emphasized in the above referenced work including an industrial system interaction denoted as the R/K trade-off, Klimpel (1984, 1987). Using a simple two parameter model for fitting time - recovery data, Klimpel (1980): [ ] where r is cumulative recovery at time t or at the t th cell, R is the long-term equilibrium recovery and K is the first-order rate of mass removal from the cell (time ), the definition of the R/K trade-off with increasing reagent dosage can be given as follows: at low dosages both R and K increase, at intermediate dosages R continues to increase but K starts to level out, and at high dosages, R may flatten out or still continue to increase but K starts to decrease, sometimes rather quickly. An illustration of this effect in a batch laboratory flotation cell is given in Figure 1. The purpose of this paper is to give a detailed plant illustration of this R/K trade-off phenomena with changing collector dosage which shows just how highly interactive the grindinglflotation system in sulfide mineral processing really is. This is one of about 20 plant level case studies that are being presented and evaluated in an upcoming book, Klimpel (1990). The particular study given here was completed in late 1979 on a copper ore of head grade 0.93% Cu. The frother being used was Dowfroth* 250 at a constant dosage of 0.15 g for each ton of solids feed to the rougher bank and the slurry pH was 10.2. The motivation of the study was to better understand the relationship of changes in feed rate (hence in particle size and state of liberation) with variations in collector dosage (potassium amyl xanthate). The major concern was to determine how much extra economic benefit could be gained by better utilizing the existing grinding mills and rougher bank consisting of eight identical cells. The remainder of the flotation circuit was also studied, including regrinding of rougher tails, but did not prove as critical as the rougher bank operation itself. It appeared from an overall economic viewpoint, the battle was being won or lost in the manner in which the rougher bank was being operated. This was not to minimize the roles of the scavenging and cleaning circuits but those portions of this particular circuit could be rather obviously operated to achieve maximum benefit given the specific output of the rougher bank. A special attempt was made to maintain constant ore type throughout the entire test program. Eight specific test conditions were evaluated including three different feed rates (and hence three different size distributions) to flotation. Tables 1 and 2 summarize the eight test conditions and overall recovery results at the end of the rougher bank for each test. A single test consisted of at least three consecutive eight-hour shifts operating at steady-state with composite cell lip samples taken over the final eight-hour shift. "Trademark of The Dow Chemical Company

Jan 1, 1990

By R. Kent Comann

Introduction This article describes ways to maximize employment potential for those looking for a job in the minerals industry. Advice on the right way to prepare a resume is followed by tips on where and how to look for a job and ways to make the best of an interview. Information on current salaries is also included. RESUME NAME: Give home and college addresses and telephone numbers, ADDRESS: if still in school. Otherwise, give home address and TELEPHONE: telephone plus business telephone number. JOB OBJECTIVE: What do you want to do? EDUCATION: List degree(s), name of college(s), location(s), and date(s) degree(s) received or to be awarded. Do not list your high school but do include postgraduate education, training, or seminars. WORK EXPERIENCE: List your most recent experience first, and the rest in reverse chronological order. Include dates (month and year) and company names. First show the entire time spent with each employer, then list the various jobs (with dates and duties) you may have had with this employer. MILITARY SERVICE: List dates, branch of service, and duties. If you did not serve, delete this section. COLLEGE ACTIVITIES: Show organizations, sports, and offices held. Only list these if you are a graduating senior, otherwise delete this section. PERSONAL DATA: List age, marital status, height, weight (optional), foreign languages spoken, professional memberships, and health (if good or excellent, otherwise do not mention it in the resume). If you prefer a certain location and would not accept a job anywhere else, list the location desired. Otherwise, do not restrict yourself. Also list special skills, hobbies, or interests that will help sell you to the company. Be brief in this section and use good judgement as to what you say. REFERENCES: Furnished on request. Resume Guidelines There are some things to remember when preparing a resume. These include: • Limit your resume to one page, unless you have a lot of experience. Then it should be a maximum of two pages. • Use 8.5 x 11 plain bond paper, white or off-white. Stay away from colored paper, odd size paper, and cute resumes. They are counter-productive. • Make certain your resume is typed professionally. Be sure to check for spelling, grammar, and typographical errors. • Have your resume reproduced by copier or printing shop. Do not use carbon copies. • Keep your sentences and remarks short and punchy. Use action verbs. • Have a good layout that is easy to read, brief, and uncluttered. Allow some white space in your resume. There are also some things you should not do when preparing a resume. These include the following: • Do not include a photograph or mention your religion, race, or sex. • Do not leave gaps of time unaccounted for in your resume. • Do not include your college transcripts with your resume or list all the courses you took in college. • Do not list all the articles you have written. Only list one or two, if appropriate for, say, a research and development job. • Do not include reference letters with your resume. Any good, prospective employer will make telephone reference checks. • Do not use "Curriculum Vitae," use "Resume." • Do not include unnecessary data such as your social security number, passport number, or PE registration number. If you are a PE or EIT, be sure to mention that in your resume. Do not mention US citizenship or US work permit unless you have a foreign education or are not a native US citizen. • Do not include present salary or salary requirements. This will come out in the interview. • Do not use the word "I" in your resume or refer to yourself in the third person. • Do not lie, distort, exaggerate, or editorialize in your resume. • Do not have your resume prepared by an outside resume service. This is unnecessary, costly, and generally not effective.

Jan 2, 1985

By Michael Chender

The rapid and continuing expansion of computer and communication capability is giving birth to profound changes in the business environment. Some commentators hold that the basis of the world economy has already shifted from industry to information. Regardless of the overall implications of these changes, it is now clear that every company potentially has both access to and a need for much more information for decision making than was the case a few years ago. The importance of information to the mining industry has been underlined by the rash of unexpected economic and political developments and by the accompanying market volatility that have marked the past decade. It should be abundantly clear to even the smallest mine operator that the price for his output can be instantly affected by unexpected developments halfway across the world. Nevertheless, many mining companies still are much less sophisticated about information than they might be. For many, the historical feeling that information is an unnecessary luxury for a production-oriented business still seems to linger. On the other hand, the overwhelming volume of source material, uncertainty as to its accuracy, and the problems of finding relevant material when needed, presents stumbling blocks to viable systems of information management. Introduction When a company needs certain information, it has two basic alternatives. It can either research and develop the information in-house or pay for someone else to do it. Few mining companies have the in-house capability to assemble information needed for careful planning. On the other hand, companies often engage consultants to find information that is in fact easily accessible. All too often companies give up on procuring needed information because of the frustration of not being able to find it, and of not wanting to pay seemingly exorbitant sums for someone else to produce it. The result is poor planning. This article addresses two aspects of the information problem-finding the right sources and then accessing the particular information desired, especially as it pertains to nonferrous metal exploration and development issues. In the first section, major English-language sources that report on exploration and development (excluding primary sources of geological information, such as technical journals) are discussed.* In the second section, some available techniques for efficiently accessing this information are presented, focusing on the new MINESEARCH database system. Part One: The Information Following the News The most obvious way to follow exploration and mining developments from a distance is through regular mining media. Many active mining districts have local newspapers, usually published in the district's major towns, that give extensive coverage to miningrelated activities. For instance, to follow activities in Idaho's Coeur d'Alene silver district, one might subscribe to the Kellogg Wardner News and/or the Wallace Miner. Similarly, one might get the Sacramento Bee for news of current mining developments in north-central California. The American Metal Market, published in New York, is the daily newspaper that covers the metals industry as a whole, both in the US and abroad. Its coverage is weighted towards the refining, manufacturing, and consumption end of the business, but it frequently carries stories on mining activity, company dealings, and market outlooks. For news of major mining companies, the daily Wall Street Journal provides occasional coverage of North American companies, while the Financial Times of London serves as probably the best single source for worldwide company information. International Coverage There are three metals industry weeklies that have broad international coverage and readership. These are Metal Week, Metal Bulletin, and Mining Journal. Each has its strengths. Metals Week, a McGraw-Hill newsletter, focuses primarily on news of the metals markets, concentrating on those developments that are seen to most directly affect the price of nonferrous metals. Metals Week is strongest on North American news. Metal Bulletin, published in London, is particularly strong on

Jan 4, 1983

By M. P. Adamson, Bryson D. Trexler, R. V. Ramani, Mark L. Koenig, V. Rajaram, Alan Burton, C. S. Crocker, Kim. Y. C., A. G. Law, William C. Larson

Another record year vas experienced by the mining industry as the value of nonfuel minerals production soared to $23.5 billion-up 15% from the previous year. Based on preliminary estimates by the Bureau of Mines, US mine output was the basis for about $225 billion worth of minerals-derived materials. Metal mining accounted for $8.5 billion, or 36%, of the total. Of the 22 metals produced in the US, 11 showed quantity increases, while 16 had value increases. Business Week's survey of 1200 companies (Mar. 17, 1980, pg. 81) revealed that 21 metals and mining companies had sales increases of 30% with earnings up 104%. Higher prices and increased demand for metals were primary factors behind the improved performance. The steel sector reported a 30% decrease in earnings for the year, but the slump was primarily the result of a $383-million loss for US Steel coupled with a 17% decrease in earnings at Inland Steel. Other firms did well in 1979, with earnings up 279% at Interlake and 269% at Kaiser Steel. Major obstacles facing the steel industry included substantial tonnages of dumped and subsidized imports, environmental expenditures, runaway inflation, and price controls. Turning to nonmetallic production, of the 49 commodities reported by the US Bureau of Mines, 29 had higher outputs and 39 had increased values. While production statistics may seem encouraging, government regulations continued to stifle industry efforts to develop needed new capacity. A case in point may be publication of the Environmental Protection Agency's simplified procedure to streamline five permit programs by consolidating them into one. The five included: hazardous waste management, national pollution discharge elimination, dredge and fill, underground injection control, and prevention of significant deterioration. The new procedure, published by EPA in the Federal Register, took 149 pages. AMC's J.A. Overton, Jr., noted that (MCJ, Nov. 1979, p. 39) to visually depict the complexity of the proposed regulations, Sell Oil Co. developed a flow chart mapping a path through one section. The chart required 10 days to map and two weeks for two draftsmen, using computerized drafting techniques, to plot the results. Overton said, "Perhaps the greatest service EPA has performed in preparing its proposal is to demonstrate again that federal regulations have become a complexity compounded by confu¬sion that culminates in a conundrum." While regulatory agencies promulgate new regulations, the lack of a cohesive government minerals policy remained a matter of significant importance to mine management. A non-fuel minerals study to provide input to the Administration for a national minerals policy program was a dismal shock, according to Rep. Jim D. Santini (D-NV). He said the first phase of the program, which was to identify industry problems, failed to come to grips with the hard decisions necessary regarding these problems, their impact, and their interrelationship. Santini said the report failed to acknowledge government's role in the process and absolved the government of any adverse influence on the domestic minerals industry. In response to the phase one report, eight senior mining executives testified before Santini's House Subcommittee on Mines and Mining concerning mineral supply, access to foreign sources, environmental protection, conservation and recycling, national security, and capital formation. By year's end, Rep. Santini promised to prepare a hard-hitting committee report on problems affecting the non-fuel minerals industry to provide a better base for subsequent policy recommendations (MCJ, Jan. 1980, p. 34). On the technological front, interest again centered on mechanization, automation, and scaleup to improve and expand production. Computer use increased for calculation of ore reserves, mine planning, operations research, and mineral processing.

Jan 5, 1980

By Richard L. Nielsen

Early Models In 1906, Daniel Jackling demonstrated that relatively low-grade disseminated copper deposits could be mined profitably using mass mining technology. The ensuing years saw a gradual increase in activity aimed at defining ore targets for large reserves of disseminated supergene enriched copper ores. Prominent organizations, such as Calumet and Hecla, Phelps Dodge, and Asarco, established teams of geologists. They focused on interpreting oxidized and leached outcrops over supergene chalcocite ore bodies. Techniques to evaluate oxidized cappings and predict the presence of underlying supergene chalcocite enrichment ores were developed and put to practical use by the late 1920s and early 1930s (Locke, 1926; Blanchard 1939, 1968). Oxidized capping evaluation centered on recognizing oxidation products derived from the various sulfide minerals found in disseminated deposits. Chalcopyrite, for example, oxidizes to a characteristic reddish pitch limonite that exhibits characteristic boxworks. Chalcocite is recognized by earthy, powdery, indigenous, hematite-rich limonite. Pyrite weathers to characteristic cavities and yellowish jarositic limonites. Evaluation of cappings is a systematic effort of noting and estimating the preoxidation sulfide mineral content and copper grade in an oxidized outcrop using mineral composition and morphology of oxidation products. Several complications were noted early in this technology's development. For example, the associated pyrite content of protore greatly influences mobility of copper and iron during oxidation. High-pyrite disseminated sulfide assemblages, upon oxidation, mobilize iron as well as copper. This results in oxidized rock with many leached cavities, poorly developed boxworks, and poorly developed indigenous limonites. This effect was noted and used in a general way to identify possible development of supergene chalcocite. Indeed, strongly leached outcrops could be present over enriched ore. A second complication is the manner in which associated gangue minerals influence the products of sulfide oxidation. Reactive gangue minerals, such as carbonates, neutralize acids formed by sulfide oxidation. Thus, copper cannot be moved from oxidized outcrops. Relatively nonreactive gangue, such as quartz-sericite-altered schist or porphyry, neutralizes supergene acid very slowly. This allows copper to move out of the oxidized outcrop. Recognizing these features was closely allied to an early understanding of the processes involved in oxidation supergene enrichment of disseminated ores. Field observations and mineralogic mapping of oxidized cappings were important in discovering many important porphyry copper ore bodies during the 1920s. These include Silver Bell, Miami-Inspiration, and Morenci in Arizona; Tyrone, NM; and Ely, NV. Capping interpretation like-wise led to discoveries at Mineral Park and Esperanza in Arizona during the exploration surge that accompanied high copper prices in the 1950s. In the 1960s, Kennecott research geologists refined and quantified some aspects of oxide capping appraisal (Anderson, 1982). The ratio of jarosite plus hematite to the total limonite assemblage was shown to be proportional to amount of copper leached from the oxidized outcrop. Thus, if the limonite assemblege in an oxidized capping consists of 40% goethite, 20% jarosite, and 40% hematite, 60% of the original copper is leached. The amount of copper remaining in the capping, together with the limonite mineralogy, can then be used to estimate original copper grade in the outcrop, amount of leaching, and chalcocite enrichment grade (in feet/% copper) in the underlying supergene chalcocite zone (Anderson, 1982). Evaluation of oxidized cappings and prediction of chalcocite blanket ore targets were successfully used in the western US and South America through the 1950s. Discovery of new porphyry copper provinces in other regions during the 1950s and 1960s, however, underscored some limitations of capping evaluation technology. Porphyry copper systems in Australia generally oxidize and weather to hematitic cappings. Presumably, the semiarid monsoon environment and the long time that the sulfide systems have been exposed to oxidation on that continent lead to development of relatively stable hematite. It is now recognized that jarosite and goethite persist metastably in the limonite assemblage in all areas once sulfide oxidation is complete and acid generation ceases (Brown, 1971). The assemblage eventually converts to hematite. But, climatic conditions in arid southwest US allows metastable limonites to persist for a long time, thus providing information on former sulfide assemblages. Another problem occurs when applying capping evaluation techniques in extremely humid re-

Jan 12, 1984

By R. W. Carn, D. Vincent

During the next 15 years, US coal production is expected to double, with the increased production evenly divided between the East and the West. Along with greater production, coal export markets should increase dramatically from East, West, and Gulf Coast ports. The annual overseas export capacity of US coal-loading terminals is expected to rise from 147.1 Mt (162.1 million st) in 1981 to a minimum of 278.1 Mt (306.6 million st) in 1985, according to the US Maritime Administration. Increased coal production and use will lead to more development of import and export terminals, a vital link in the coal transportation chain. With continually escalating capital costs and the competitive markets that the terminals will serve, a well designed and efficient terminal is necessary. This article begins a two-part series that presents concepts presently used in coal export terminal design. Part I looks at site selection factors and equipment needs, while Part II will examine environmental considerations in building a terminal as well as typical capital and operating costs. The world is nearing the end of the oil era. In a few years oil will not be available to sustain the growth rate and increasing standard of living we have known in our lifetime. The big question is what energy era are we moving into? With the decline of readily available oil reserves and rapidly increasing prices, many countries are trying to switch to alternate energy forms. While intensive efforts to find new oil reserves continue, alternate energy sources such as natural gas, coal, synthetic fuels, nuclear, hydroelectric, solar, and wind power are being developed. Recent indications are that coal is expected to bridge the energy gap over the next 25-30 years until the technology and economics of the alternate energy forms reach satisfactory levels. Use of coal for energy is receiving strong attention due to its long-term availability (200-300 years minimum), relative ease of development, and its low cost per unit of power produced. By the year 2000, it is expected that 25% of world energy supply will be met by direct coal combustion and possibly another 5-10% by synthetic fuel from coal. Coal's expanding share in the world energy market, along with an increase in coking coal requirements, will result in a large increase in the world's seaborne coal trade. Recent statistics and projections for the future are shown in Table 1. This phenomenal development rate includes increases in both coking and thermal coal requirements. Because of the rapid increase of seaborne coal trade during the last 10 years and the even greater projected increase of trade to 2000, various sectors of the coal industry are faced with enormous technical challenges and huge investments in equipment, land, transportation systems, and port facilities. Very large bulk terminals are under development throughout the world. Latest surveys indicate that there are about 30 new coal export and import terminals under consideration and at least 30 existing terminals have expansion programs planned or underway. With the high cost of borrowed capital and rapid inflation rates there is great emphasis on new planning and design techniques to minimize capital and operating costs of coal transportation systems. Terminals A total coal supply system can be considered to consist of one or more mines; a train, barge, truck, or other haulage system; an export terminal; a fleet of bulk carriers; a receiving terminal; and possibly, local inland distribution networks that include barges and railways. Terminals, though only a small link in the total transportation system, play a key role in overall system efficiency. At ports or inland distribution centers, terminals act as transportation links bringing trains, ships, barges, or trucks together for cargo transfer and temporary storage. A well-designed terminal can provide maximum independence between two modes of transportation and optimum freedom for intermodal interference. A terminal acts as a buffer between the two transportation modes by providing sufficient storage capacity so a ship need not wait for its cargo on, for example, a train-by-train basis, but can load immediately from the ready stock. Similarly, a train need not wait for a ship to unload its contents but can dump immediately into storage. A terminal also can be used to properly mix various types of coal to satisfy a buyer's requirements. Consider the relative value of various production and transport segments for a typical steam coal

Jan 6, 1983

By M. E. Rossi

Reconciliation of long-term, block-model reserves and mine production frequently shows significant discrepancies. This reconciliation can be done comparing the long-term, block-model reserves against a grade control model (a model based on blast holes or production data). In this article, this is called an Fl factor for tonnages, grade and ounces. This reconciliation should also be done by comparing the grade control model to the mill feed (or loaded to heaps) tonnages, grades and ounces (F2 factor). The long-term model usually contains blocks with dimensions larger than the size of the selective mining unit (SMU) of the operation. The smoothing effect of an estimator, such as kriging, will generally result in a grade-tonnage curve that does not match the grade-tonnage curve of the SMUs to be recovered. This problem is partly to blame for the reconciliation discrepancies found. In addition, in-pit perfect selection is impossible. Therefore, the grade-tonnage predictions based on blastholes also need to be assessed - considering unplanned dilution and the errors of estimation of the grade control model. This article focuses on using conditional simulations to assess the uncertainty of the predicted grade-tonnage curve in terms of minable reserves and in the framework of long-term mine planning. Quantifying the uncertainty in no uncertain terms is done in conjunction with a popular pit optimizer. It yields several possible minable reserves and several possible minable tonnages and grades, per pit shell or scheduled phase. This is a step further from the usual sensitivity analyses that are performed for deposit evaluation or development planning. Specifically, this approach takes into account the volume-variance relationship, the selectivity of the mining operation and the planned and unplanned dilution. It also carries the local uncertainty implicit in grade estimation into the mine design and mine planning phases. The estimation of recoverable reserves revolves around estimating the block distribution of metal grades from the drill hole information available, typically composites of a fixed length. Ideally, this estimation should accurately predict the ore sent to the processing facilities of an operation. Here are some definitions used in this article: ? Recoverable reserves are those where an attempt has been made to estimate the actual distribution of grades (usually by a block model) that would be found in the ground at the time of mining, relative to the mine equipment size and other operational characteristics of the mine SMU. In particular, it incorporates the effect of internal dilution for each block how the grade distribution changes in variability when comparing a composite grade distribution to the "mining unit" grade distribution. In geostatistical jargon, this is known as "change of support effect," or volume-variance effect. ? Minable reserves are those recoverable reserves defined by a mine plan. In an open-pit scenario, they would incorporate pit optimization in addition to the sequencing and scheduling of production through mine life. ? All other block model estimates from drill hole data that do not explicitly attempt to estimate or measure the volume-variance effect are called in situ. It can be argued that these models may be inappropriate tools for predicting the grade-tonnage curve of the material to be sent for processing (minable reserves). The purpose of estimating recoverable reserves can be thought of as estimating the appropriate amount of dilution that should be incorporated into the model so that it adequately predicts (long-term and on average) the grade-tonnage curve of the ore sent to the processing facilities (mean grades and ton-

Jan 1, 1999

By Kelly F. Bailey

Introduction On June 20, 1986, the Occupational Safety and Health Administration (OSHA) published revised asbestos exposure standards for general industry and construction. The standards reflect OSHA's attempt to adequately control workplace exposures to minerals it considers carcinogenic - minerals capable of causing or contributing to cancer. These standards specifically identify asbestos as: chrysotile, an asbestiform serpentine mineral; and the amphibole minerals amosite, crocidolite, tremolite asbestos, actinolite asbestos, and anthophyllite asbestos. Each of these has a more common nonasbestos mineral analog that exists in nature in a crystalline, blocky shape rather than the hair-like or fibrous shape of asbestos. The mineralogical names for three of these nonasbestos minerals are unique: antigorite for chrysotile, cummingtonite-grunerite for amosite, and riebeckite for crocidolite. The other three nonasbestos analogs do not have unique mineralogical names. They are simply designated as actinolite, tremolite, and anthophyllite without the word asbestos following their names. The 1986 OSHA standards not only cover exposure to the six asbestos minerals, they also cover specifically the nonasbestos forms of actinolite, tremolite, and anthophyllite (AT&A). The new standards regulate these minerals exactly like asbestos (OSHA, 1986). The construction aggregate industry views this as a major problem because these nonasbestos minerals are common amphibole rock-forming minerals in the earth's crust. They exist in small quantities over large areas of the United States (Kuryvial et al., 1974). These minerals, unlike asbestos, are not mined for a specific commercial purpose. They are unavoidable components in much of the aggregate used for construction throughout the US. They are also common in the gangue material of metallic ores. There are areas of the US where amphibole-bearing bed¬rock is common. Not every rock mass in these areas contain amphiboles, however. It does mean, though, that amphiboles are physically compatible with many of the rocks in those areas. And given the correct geochemical conditions, they will be present primarily in the nonasbestiform variety. In addition, these amphiboles will probably exist in the natural drainage system, sand and gravel deposits, stream sediments, lake shores, valley basins, or ordinary beach sand within these areas. There has been little quantification of nonasbestiform AT&A in dusts and soils in the US. This is not surprising since these nonasbestiform minerals are not commercially valuable. However, an example of the pervasive nature of these minerals can be found in a 1981 Geological Society of America publication where about 0.7% tremolite-actinolite was found in the desert dust in and around Tempe, AZ (Pewe, 1981). Since OSHA standards treat these common nonasbestos minerals as carcinogens in the same way as asbestos, large natural areas in the US are implicitly being labeled as hazardous by OSHA. When a substance is identified as a carcinogen, another OSHA standard comes into play, the Hazard Communication standard. There are also right-to-know laws in 9 states that essentially duplicate this federal standard. These standards require that a product containing 0.1% or more of an OSHA-designated carcinogen be labeled as such (OSHA, 1983). This means that much of the stone and sand gravel products occurring naturally and mined in the US could be labeled a carcinogen when, in fact, they are not. The National Stone Association (NSA) and the domestic construction and mining industries believe that OSHA has seriously erred. The NSA has studied the health, mineralogical, technical, economic, and legal basis for OSHA's action. These studies concluded that there is no justification for the agency regulating nonasbestos minerals as if they were asbestos. Health issues The preamble to OSHA's 1986 asbestos standard states that evidence for asbestos-like health effects from exposure to nonasbestiform varieties of AT&A is inconclusive (OSHA, 1986). The fact is, not only are the data inconclusive, they are nonexistent. During 1986-1987, NSA's occupational health and epidemiology consultant, Environmental Health Associates (EHA), reviewed all available health studies related to AT&A. EHA found evidence that malignancies in both experimental animals and humans are associated with the asbestos forms of these minerals. No experimental or epidemiological evidence was found that indicated such pathogenic effects occur from exposure to nonasbestiform varieties of these minerals. There are relatively few scientific studies of the health effects of exposure to nonasbestiform varieties of AT&A. In three different animal studies, exposure to either nonasbestiform tremolite or actinolite did not result in pulmonary fibrosis on in excess tu-

Jan 11, 1988

By E. Haug, Quamrul Ahsan, R. Norcross

Cyanide leaching is the most common process employed in the majority of the gold extraction plants. These gold extraction plants naturally generate tailings and/or effluents that contain highly toxic cyanides, In geographic locations with excess precipitation and therefore net positive water balance, a portion of these tailings must be discharged. Before these highly toxic cyanide-containing effluents can be discharged to the waterways, cyanide and any other toxic substance must be destroyed. Geographic locations with high evaporation and low precipitation, as in the western USA, result in net negative water balance and therefore these plants can operate with a zero discharge option. However, with the current stringent environ­mental regulations and concerns for bird and wildlife mortality, even these operations may be required to partially treat their tailings to preclude potential harm to wildlife. Furthermore, heap leach opera­tions are now generally required to detoxify the heaps before abandonment. It is, there­fore, apparent that all gold extraction plants may be required to install some form of effluent treatment in order to achieve the above objectives,.

Jan 1, 1990

By Keivan PakIman, Maziar Pasdarpour

Transportation tunnels are highly sensitive to ground movements and are required to be stable and secure for safe operation. In urban areas, the interaction between tunnels and adjacent piles, which are widely used to support buildings and other structures, requires careful consideration. The seismic response of tunnels may vary depending on the lateral stiffness and strength of the tunnel walls, while the axial loading and seismic impact of adjacent piles may cause soil displacement and stress changes that can damage the tunnel lining. As a result, tunnel owners often impose strict limits on the design of new piles lying close to existing tunnels including maximum allowable tunnel deformation, stress change limits, and minimum tunnel to pile separation distances. This paper describes a numerical analysis of the inuence of the axial loading and seismic impact of bored pile foundations on subway tunnels, undertaen using the nite element software laxis D. ased on the results of the numerical analysis, the paper suggests a more realistic and less conservative approach for designing piles in close proximity to transportation tunnels. INTRODUCTION The interaction between tunnels and adjacent piles can cause stress changes and displacements in the soil and structures which may adversely impact structural performance and tunnel safety. It is therefore important to take account of relative load transfer and piling conguration when designing new piles in

Jun 23, 2024

By Igor V. Nazarov

INTRODUCTION Tens of deposits in' coal basins of Western and Eastern Siberia are being open-mined with the technology of overburden rehandling into the mined out space of pits by excavators - draglines. This technology is preferred due to the high productive Russian equipment, comparatively low cost, and relatively small Influence exerted on the environment. The basic tendency of the last decades is the aspiration for enlarging the sphere of the' technology application for development of gentle and steeply dipping seams with the use of blast destruction of overburden and with combined winning of loose alluviums and rock partings. Complex technological schemes with multiple rehandling are widely used. In addition to it, stripping workings, intermediate storage of mineral resources, placing of transport dumps, and engineering construction can be realized with the use of draglines. More than ten of dragline faces are simultaneously exploited in some pits, and the rehandling coefficient ranks from 0.1 to 8. Here, conditions of winning and placing (position of a face and existing dump) are varying in space and time; this leads to the necessity of reconsidering the order for conduction of works (schemes) within the current designing. By now, in order to this, it is required to realize complex engineering calculations by graph-analytical methods and definite experience. The need is available for automatization of this process by creation of the corresponding software. The software is to: • Orient on the active pits and use information on the existing geology, equipment, and situation in mine. • Calculate the existing schemes more operative and on a higher qualitative level to model their development in a pit space with parameter optimization. • Predict situations demanding to change the existing technology for mining operations and recommend optimal technological schemes. METHODOLOGICAL FOUNDATIONS FOR MODELING The problem is not new, it has more than fifty year-old history of creating mathematical models algorithms, and programs in Russia, Ukraine, USA, Canada. Australia and other countries with the developed mining branch. Conventionally, all population of programs can be divided into three groups depending on goal and interval of prediction: • Long-term programs used in designing and intended for substantiation of the nine capacity, selection of mining system, establishment of limits for applying the technology of overburden rehandling. Mid-term programs used in current planning for calculation of technological schemes and certificates of face in active mines.INTRODUCTION Tens of deposits in' coal basins of Western and Eastern Siberia are being open-mined with the technology of overburden rehandling into the mined out space of pits by excavators - draglines. This technology is preferred due to the high productive Russian equipment, comparatively low cost, and relatively small Influence exerted on the environment. The basic tendency of the last decades is the aspiration for enlarging the sphere of the' technology application for development of gentle and steeply dipping seams with the use of blast destruction of overburden and with combined winning of loose alluviums and rock partings. Complex technological schemes with multiple rehandling are widely used. In addition to it, stripping workings, intermediate storage of mineral resources, placing of transport dumps, and engineering construction can be realized with the use of draglines. More than ten of dragline faces are simultaneously exploited in some pits, and the rehandling coefficient ranks from 0.1 to 8. Here, conditions of winning and placing (position of a face and existing dump) are varying in space and time; this leads to the necessity of reconsidering the order for conduction of works (schemes) within the current designing. By now, in order to this, it is required to realize complex engineering calculations by graph-analytical methods and definite experience. The need is available for automatization of this process by creation of the corresponding software. The software is to: • Orient on the active pits and use information on the existing geology, equipment, and situation in mine. • Calculate the existing schemes more operative and on a higher qualitative level to model their development in a pit space with parameter optimization. • Predict situations demanding to change the existing technology for mining operations and recommend optimal technological schemes. METHODOLOGICAL FOUNDATIONS FOR MODELING The problem is not new, it has more than fifty year-old history of creating mathematical models algorithms, and programs in Russia, Ukraine, USA, Canada. Australia and other countries with the developed mining branch. Conventionally, all population of programs can be divided into three groups depending on goal and interval of prediction: • Long-term programs used in designing and intended for substantiation of the nine capacity, selection of mining system, establishment of limits for applying the technology of overburden rehandling. • Mid-term programs used in current planning for calculation of technological schemes and certificates of face in active mines.

Jan 1, 2002

By E. Wiewiorowski

A new technology for recovery of metals from spent catalysts containing aluminum, cobalt, molybdenum, nickel and vanadium was developed. The catalysts are decomposed in two consecutive pressure leaching steps. Sodium aluminate-sodium hydroxide systems are lixiviants in both leaching steps. Selective extraction of molybdenum and vanadium is achieved in the first leaching step. The leach liquor is processed for separation and recovery of these two metals. The residue from the first leaching step is further processed in a second leaching step where alumina is solubilized. The residue from this operation is a nickel-cobalt concentrate. Alumina is separated from the second stage leach liquor by precipitation; the depleted liquor, a sodium aluminate solution, is recycled to the first and second leaching step. This technology has advantages over other approaches for the recovery of metals from spent catalysts because no solid waste is generated.

Jan 1, 1988

By D. J. Brunner, S. Mathur, D. McKinney

With the advent of faster micro-processors, the use of numerical methods to simulate complex fluid dynamic phenomena in three dimensions for use in design has become prevalent in the automotive, and turbo-machinery industries. The Computational Fluid Dynamics (CFD) method divides the region of interest into small control volumes which form the mesh representing the physical characteristics of the problem, and uses the finite volume method to intergrate the equations for the conservation of mass, momentum, energy and species over each control volume. Recent developments in CFD software expedite mesh generation, and enable the use of unstructured grids, comprised of tetrahedral volumes in three dimensions and triangular areas in two. CFD more accurately represents complex geometries and allows for relative movement of meshes enabling simulation of multiple moving bodies. 'ibis paper presents two examples of how CFD simulation can be used to assess mine and tunnel ventilation problems formerly addressed by application of analytical solutions which were developed assuming ideal incompressible conditions. CFD simulation is used to evaluate the impact of varying the airflow in a descentionally ventilated airway on the layering along the roof of smoke and hot gases resulting from a vehicle fire. Control of the smoke layer is required to enable safe egress from the vehicle, particularly if the vehicle is for personnel transport, and to ensure control of the fire contaminants throughout the ventilation system. The airflow required to prevent layering against the ventilation direction, calculated from the Bakke and Leach relations (Bakke and Leach, 1962), is compared with the CFD simulation results. An evaluation of the pressures, generated as a vehicle enters a tunnel portal, using CFD simulation, is also presented for unflared and flared portal configurations. These simulation results are compared with predictions derived using an analytical method which assumes one-dimensional and incompressible flow. Results of the CFD simulation are presented in an animated video format. SIMULATION OF BACKLAYERING In designing a ventilation system for a transit tunnel, the ability of the ventilating air to control and prevent backlayering of smoke and hot gases resulting from a vehicle fire is of prime concern. The buoyant nature of hot smoke causes it to rise relative to the colder, fresh air provided by the ventilation system. If the vehicle fire occurs in a descentionally ventilated tunnel, the smoke may tend to move upgrade in a layer above the incoming ventilation airflow. The layer may become thick enough to engulf a substatntial part of the tunnel cross-section upgrade of the incident that comprises the evacuation route. This effect is termed "backlayering' and it is similar to the development of methane layers in mines for which most studies related to backlayering have been done. Prediction Techniques Analytical A number of studies have been conducted (Bakke and Leach, 1962) to define the characteristics of this phenomena and as a result have produced relations which are used both in the mine and transit ventilation fields to define the air velocities required to control layering. In the transit industry the air velocity required to prevent the backlayering phenomena from occuring during a vehicle fire is called the "critical velocity" (Associated Engineers, 1975) and is dependent upon a number of factors: tunnel height, cross-sectional area and grade; ambient air temperature and density; and the heat release rate of the fire. Common practice in transit ventilation design is to provide an airflow which meets or exceeds the critical velocity. In order to determine whether or not the critical velocity can be achieved with a particular ventilation system, a one-dimensional simulation of the tunnel network is typically performed using programs such as the Subway Environment Simulation program (SES) originally developed in the late 1970's (Associated Engineers, 1980). The results obtained from SES are compared to the critical velocity to determine the adequacy of the ventilation system. Computational Fluid Dynamics For the backlayering simulations, a commercial CFD code which has been used successfully in a wide variety of engineering applications, was used. It provides numerous options for modeling laminar and turbulent flows, multiple turbulence models, definition of multiple species and chemical reactions between them, a variety of boundary conditions (including constant pressure and constant velocity inlets) and the ability to apply user-defined FORTRAN subroutines. It includes the ability to model conductive, convective, and radiative heat transfer. FLUENT also permits the use of "body-fitted coordinates" to match the computational mesh or grid to complex real-world geometries. Computational Fluid Dynamics Model The model developed to simulate the backlayering phenomena is comprised of an airway of rectangular cross-section, 4 meters wide, 4.5 meters high, and 200 meters long. A laterally

Jan 1, 1995

By L. Micale

In May of 1986, the Arizona Legislature passed comprehensive environ-mental legislation called the Environmental Quality Act (the Act). The Act focuses on ground water protection and addresses drinking water quality, aquifer protection standards, requirements to apply technology to reduce discharges, and a state Superfund program.

Jan 1, 1988

By R. K. Singhal

In open pit mining every effort is made to maintain as steep a pit slope as possible. Pi t slope angles are influenced by numerous factors some natural that are amenable to limited or no control at all and other mining technology related where a greater degree of influence can be exerted by the engineer. One such area where control is possible is the method of drilling and blasting in the pit. This paper discusses the influence of blasting techniques on the economics of pit slopes and demonstrates how balanced blasting can be used with advantage to pit wall control and improve fragmentation (thus reducing material handling costs). Microprocessor based instrumentation to monitor drill and blast performance has also been described.

Jan 1, 1988

By D. R. Barker

The economic performance and competitiveness of a concentrator are dependent on the production of accurate real-time assay information and the manner in which it is utilized. Our paper describes the relative merits of on-and in-stream analysis techniques and the economic significance of Assay-Based control and Target Oriented Display Systems. At Outokumpu's pyhasalmi Concentrator, a Flotation Expert Display System aids the process operator to make better decisions and provides guidance on the most efficient method of achieving the desired control objectives.

Jan 1, 1989