Search Documents

Objective To identify techniques to reduce the probability of gob gas explosions. Background In the last 6 years, seven explosions of methane and/or coal dust occurred within worked-out, sealed areas (gobs) of underground U.S. coal mines. These explosions, believed to have been started by lightning, destroyed many mine seals and caused much damage external to the sealed area. Three of these gob gas explosions occurred in one mine in Alabama over a 3-year period. Fortunately, the explosion forces and the toxic gases that vented from the sealed gob area did not cause fatalities or injuries, but they destroyed several large seals. If miners had been inspecting or working near these areas, the potential for serious injury and/or death would have been high. The United Mine Workers of America and the Mine Safety and Health Administration (MSHA) requested that the National Institute for Occupational Safety and Health (NIOSH) help identify the mechanisms for lightning penetration into the gob and recommend ways to reduce the probability of occurrence of future explosions from such lightning penetrations into sealed areas of underground coal mines. Because three of the explosions occurred in a single mine, special attention was focused on these particular explosions. The first occurred in April 1994 in a sealed area, which enclosed about 1.35square miles of abandoned workings (gob). This explosion destroyed 3 of the 38 seals that surrounded the gob. These seals were less than 2,000 ft away from three 4.5-in-diam steel-cased test wells that extended from the surface into the mine entry. At the time of the explosion, the National Lightning Detection Network (NLDN)verified 12 lightning strikes within 10 miles of the mine, including several above the gob. After the 1994 explosion, the damaged seals were rebuilt using an acceptable seal design capable of with standing a 20-psi explosion, as required by 30 CFR 75.335. On January 26,1996, a second gob explosion destroyed five more cementitious pumpable seals less than 2,000 ft away from those destroyed in 1994and even closer to the steel-cased wells. The NLDN verified72 lightning strikes in the area of this second gob explosion. Compressive strength analyses of fragments from the destroyed seals showed that strengths ranged from 11 to 138 psi, with an average of83 psi. This is over 100 psi below the minimum 200-psi compressive strength requirement for cementitious pumpable seals. These seals were again replaced. On July 9, 1997, the third and most violent explosion occurred in the same vicinity of the last two explosions. Three more cementitious pumpable seals were destroyed, including one newer seal that exceeded the minimum 200-psi compressive strength requirement. The NLDN verified 695 lightning strikes above the mine during the time in which the explosion occurred. MSHA's

May 1, 2001

By L. Yacher

Most metallurgical plants already have sophisticated control systems as well as a process-historian that manage thousands of real-time and historical variables. There exists a huge potential for operational improvement through the inference of patterns, correlations and cause-effect relationships within these records. The ability to run online and offline advanced multivariate statistical analyses represents a great potential for the mining industry, in various areas such as cathode quality improvement in an SX/EX plant, the analysis of the correlation between the ore characteristics and the mill throughput, the determination of early failure alerts for major equipment, etc. The background theory is described, as well as some practical experiences of operations in Chile and Peru. Cu2007

Jan 1, 2007

By R. L. Marek

The number of persons fatally injured in the petroleum industry in California in 1930 was 50, a decrease of 20.6 per cent from 1929, when 63 men accidentally lost their lives. The decrease in the number of fatalities for 1930 has caused the first downward trend in the fatality curve of the California petroleum industry since 1927 and is coincident with a period of little active development in new fields, the shutting in of many wells, and reduced activity in the older fields of the State. That a direct relationship exists between the number of accidental deaths and the rapidity of new field development is particularly emphasized by the fact that in 1929, a year of intensive field development in California, 48 men were killed in the drilling and producing division, whereas in 1930, a year of less development work, only 25 men were killed in this division. The figures for the average number of drilling and producing wells in 1930 decreased appreciably from, comparable figures for 1929 and a greater percentage of producing wells were shut-in in 1930 than in 1929.

Jan 1, 1931

For the year ending February, 1918 PRESIDENT PHILIP N. MOORE, ST. Louis, Mo. PAST 'PRESIDENTS WILLIAM L. SAUNDERS NEW YORK,. N. Y. L. D. RICKETTS NEW YORK, N. Y. FIRST VICE-PRESIDENT SIDNEY J. JENNINGS NEW YORK, N. Y. TREASURER GEORGE C. STONE, NEW YORK, N. Y. SECRETARY EMERITUS ROSSITER W. RAYMOND, _ NEW YORK, N. Y. SECRETARY BRADLEY STOUGHTON NEW YORK, N. Y.

Jan 11, 1917

By R. L. King

An expert system is a computer system that performs tasks that are considered by most people to require human expertise. Since their functioning relies heavily on large bodies of knowledge, they are often referred to as knowledge-based systems. These systems can contain the skill and knowledge of an individual human expert or the combined knowledge of a group of experts. The knowledge embodied in an expert system can consist of specific facts, HSHA regulations, heuristics or rules-of-thumb used by the expert, and databases. The purpose of this paper will be to introduce the participants in the ?Increase Competitiveness Thru Innovation? session to expert systems. This introduction will enable the participants to better appreciate the balance of the papers to be presented in this session. These papers will deal with specific mining problems and a proposed innovative solution. This paper will discuss such topics as problem-solving paradigms, rules of inference, knowledge representation, and the structure of expert systems.

Jan 1, 1989

By George Weinman

GEORGE WEINMAN: Our first group leader for the discussion, "Ups and Downs of Downsizing," wishes to remain anonymous. ANONYMOUS PRESENTER: When George and Jamie asked me to do this, I thought a while about what to say and went through several variations of what I’ll present today. I came up with the title "Ups and Downs of Downsizing." A lot of downsizing has occurred in all of our industries-mining in particular. While this downsizing has gone on there's no doubt that all of us produced more and more, probably with less and less people and less and less resources. But I contend that there's also a big downside associated with this. I disagree with a point made in the keynote presentation yesterday, the point being that vendors are bringing new technology to us. I can agree that the major changes associated with equipment and similar items are being brought to us by vendors. But I do contend that there's an awful lot that's getting away from us-especially if you factor in the integrated supply or the strategic alliance or the various and sundry related purchasing techniques. I contend that we are missing a lot of, let's call it, lesser technology; not major equipment changes, but different tools, different techniques for doing things, different left-handed widgets. The integrated suppliers, to some degree, are bringing this to us; but I am concerned about the small guy who comes up with a good idea but deals through integrated suppliers rather than having direct contact with plant people. They're probably going to go somewhere else and market their wares. That's what I'm hearing from a lot of smaller vendors. I also think that we are losing a lot of our knowledge base. Every time you shrink an organization somebody leaves. The bottom line: less people. When a person leaves-and I don't care who he or she is, be it the janitor who's cleaning the bathroom, the operator, the engineer-when that person leaves there's some knowledge that goes with that per- son. And sometimes that knowledge is quite valuable. I've worked more or less at the plant "low." I can't tell you how many excellent, viable projects have come from people who said, "Look here. There's a better way of doing this." And, you know, they were right. Now, maybe they didn't have the resources or the expertise to develop these ideas into full projects, but they had that basic idea. Someone may say, "Well, if we put a no. 6 skyhook here, so we can lift this particular pump without bringing a crane from 30 miles away ...." No. 6 skyhooks are a bit expensive, but what if we put together a little structural steel here or whatnot?

Jan 1, 1998

By P. L. De Bruyn, R. J. Charles

The transfer of kinetic energy of translation into other forms of energy by impact is a fundamental process in most crushing and grinding operations. During and after the impact process the original source energy may be accounted for in any of the following possible forms: 1) Kinetic energy of translation of both the impacted and impacting objects. 2) Kinetic energy of vibration of the components of the impact system. 3) Potential energy as strain energy of the components of the system or in the form of residual stresses. 4) Heat generated by internal friction during plastic deformation or during damping of elastic waves. 5) New surface energy of fractured materials. At any instant during the impact process only the strain energy of the components of the system can contribute directly to the brittle fracture process. If fracture is the desired result, as in comminution, it would seem advantageous to choose or arrange the conditions of impact so that a maximum amount of the original kinetic energy could be converted to strain energy at some moment during a single impact. The present work deals with determination of these desirable conditions for a simple case of impact and application of the principles involved to general cases of impact.

Jan 1, 1956

By H. L. Hartman

The estimation of the magnitude of pressure changes which occur in mine ventilation circuits is of primary importance to the mining engineer in making changes in an existing mine or in projecting the ventilation requirements of a new mine. It is a formidable task in many cases, because no adequate theory is available and empirical data are unreliable or even lacking. The common practice of increasing one's estimate of the mine static pressure by an arbitrary "safety factor" to compensate for the unknown is a poor substitute for accuracy. Generally little difficulty is encountered in calculating friction losses in mine airways. As long as the friction factor can be reliably estimated, the well-known Atkinson formula gives accurate results. However, it is in the determination of shock losses that the ventilation engineer encounters difficulty and is apt to resort to an arbitrary allowance. Shock losses may constitute as little as five or as much as twenty-five percent or more of the overall pressure drop in the mine, and such allowances are generally completely unsatisfactory. Since other shock losses have been covered by McElroy (1), this paper deals with a long-neglected source of shock and pressure change in ventilation circuits, divergence and convergence of airflows, here referred to by their more common mining terms, splits and junctions. PREVIOUS WORK There is practically no fundamental information available in ventilation literature on the splitting and junction of airflows, either in mining or mechanical engineering periodicals or handbooks. Fluid mechanics textbooks seemingly avoid the subject of divergence and convergence, except for brief empirical treatment. The only investigation reported for mine ventilation systems was by Weeks, et a1 (21), in 1933. Although applicable to the case of line brattices in coal mines, it cannot be applied to the general case of mine splitting and junctions involving change of direction with little or no area change. In dealing briefly with the subject of shock losses at splits and junctions, McElroy (1) suggests only that they be considered comparable to bends, with losses at junctions being increased 50 pct to allow for interference of merging streams. He recognizes the importance of quantity ratio but concludes that exact quantitative data are lacking to compute losses at splits and junctions. A very capable summary and analysis of the state of knowledge in industrial ventilation and fluid mechanics regarding losses due to divergence has recently been reported by Gilman (3). In considering all earlier work dealing with both air and water as fluids, he compares results on a common basis and obtains good empirical agreement. For divergent flow, the controlling variable in each branch is the quantity ratio, although the deflection angle is of secondary importance. Empirical equations are presented for straight and 90-deg branches with varying velocity ratios, assuming branches smaller than the main duct. These formulas are modifications of the Borda equation for abrupt contraction and expansion. The only information available on shock losses in convergent flow are approximate experimental data presented by Alden (4) and the Manual of Recommended Practice for Industrial Ventilation of the American Conference of Governmental Industrial Hygienists. Alden considers the effects of both quantity ratio and deflection angle. The data indicate an effect opposite to that observed in divergence, the shock loss varying directly with quantity ratio. SHOCK LOSS THEORY Because pressure losses due to shock have been found to bear a constant relation to the mean velocity of flow in a given conduit, they are frequently expressed as a dimensionless function of the velocity head, termed the shock loss factor. For airflow in ducts and mine airways, the shock loss may be represented by the formula (1): Hx - XHv, (1)

Jan 1, 1961

In the interest of safety in metal mining, the Bureau of Mines recommends: That underground metal mines be equipped with a main fan in a fireproof housing, located preferably on the surface. The fan installation should be designed to ventilate the mine adequately throughout and to permit prompt reversal of direction of air flow. Many metal mines depend for ventilation upon natural flow of air, assisted by the exhaust from pneumatic drills and by compressed air blown into the workings. Others also ventilate separate headings or sections with fans and blowers. Relatively few have o mechanical ventilating system that controls the coursing of air throughout the mine.

Jan 1, 1939

By J. W. Powell, F. H. Spedding

A complete review of the use of chelating agents in the sepa ration of rare earths by ion-exchange is given as well as a concise description of the recent pilot-plant operations of the Ames Laboratory. The two chelating agents which show the greatest promise are ethylenediamine-N,N,N',N'-tetraacetic acid and N'-(2-hydroxyethyl)ethylenediarnine-N,N,N '-triacetic acid. The first successful separations of rare earths by ion exchange were reported in a collection of papers which appeared in the November, 1947, issue of the Journal of the American Chemical Society.1'9 Some of this work was performed at the Ames Laboratory of the A.E.C., the remainder at the Oak Ridge National Laboratory. The processes developed at Oak Ridge, as well as some of the early Arnes methods, employed 5 pct citric acid-ammonium citrate eluant at low pH and were carried out on either H+-state or NH:-state resin beds. These techniques were successful for the separation of small quantities of either naturally occurring or radioactive rare earths and are still used for the isolation of rare-earth activities from fission products. Concentrated citrate is not economical, however, for use in moderate or large-scale rare-earth separations. For this reason, the Ames Laboratory turned its attention to lower concentrations and higher pH's in order to make more effective use of the eluting agent.10-14 Although elutions have been performed successfully over a wide range of conditions, 0.1 pct citrate at a pH of 8.0 is highly recommended for use on H+-state resin beds.15,16 Elu-tion with 0.1 pct citrate in the pH range from 5 to 9 brings about the separation of the constituent rare earths into a series of flat-topped elution bands which progress down the resin bed, head to tail, without actually pulling apart as do the rare-earth peaks which develop when trace quantities are eluted with 0.25M (5 pct) citrate at low pH's. Because of this, elution with 0.1 pct citrate at pH 8.0 do not produce pure rare earths unless sufficient quantities are present to provide developed bands which are at least several inches long on the columns. Although other eluting agents have proved more effective than citrate solutions, articles concerning the use of citric acid for separating rare earths still appear in the literature occasionally. For example, Ketelle and Boyd17 reported some further studies on the separation of rare earths with 5 pct citrate in 1951. They used 270 to 325 mesh Dowex-50 columns at 100°C and a pH of 3.28. vickery18 compared the effectiveness of a number of eluants for the separation of rare earths on NH+4 state Dowex-50 in 1952. He found that citric acid was more efficient than acetic, malic, tartaric, and aminoacetic acids. In 1953, Mayer and Freiling18 reported that citrate was inferior to malate, glycolate, lactate, and EDTA for the resolution of Sm-Eu and Y-Tb mixtures on 250 to 500 mesh, NH+4-state, Dowex-50 at 87°C. pinta20 used 5 pct citrate in the pH range 2.8 to 3.4 to obtain some rare earths for analytical work the same year. Trombe and Loriers21 reported the use of citric acid in their laboratory to separate rare earths in kilogram quantities. Lariers and Quesney22 reported some separations with citrate in 1954. They used 5 pct citrate at pH 2.8 to separate yttrium and the yttrium-group rare earths from the cerium-group rare earths. briers23 used 5 pct citrate at pH 3.2 and 90°C as late as 1956 to isolate thulium in fair purity. For the separation of tracer quantities of rare earths, Mayer and Freiling19 have recommended pH 5.00, 0.24M lactate at 87 °C on 250 to 500 mesh, NH+4-form Dowex-50. Freiling and Bunney24 have also employed lactic acid for the separation of fission-product rare earths. Cunninghame, Size-land, Willis, Eakins, and Mercer25 have reported a 4-hr separation of Y, Eu, Sm, Pm, Nd, and Pr with 1M lactic acid at pH 3.25 on Zeokarb-225 at 87°C. Stewart, et al.,26 reported separation factors for rare earths distributing between Dowex-50 and 0.25M glycolic acid. stewart27 also reported a 30-61 separation of Y, Tm, Er, Tb, and Lu tracers with buffered, 0.25M glycolic acid containing 0.05 pct Aerosol OT. The resin bed was 400-mesh Dowex-50 (X12). The pH of the ammonia-buffered eluant was 3.5. Various aminopolyacetic acids have also been used to obtain varying degrees of separation of the rare earths. In 1951, Fitch and Russell28 investigated iminodiacetic acid (IDA) and nitrilotriacetic

Jan 1, 1960

By D. M. Maijer, S. L. Cockcroft, A. B. Phillion, P. Fan, C. Reilly, H. Khadivinassab

A combined experimental / numerical approach is proposed to investigate the bulk transfer of solute due to liquid metal feeding during shape casting of aluminum alloy A356 (Al-7Si-0.3Mg). A dumbbell-shaped experimental casting geometry has been developed, which promotes solute redistribution due to liquid metal feeding. The redistribution of solute in trial castings is evaluated based on the area fraction of eutectic phases. In the numerical model, silicon species segregation during solidification are calculated assuming the Scheil approximation, and are coupled with macro-scale solute transport due to Darcy flow in the mushy zone and feeding flow. The model has been implemented within the commercial CFD software, FLUENT, which simultaneously solves the thermal, fluid flow fields and species segregation on the macro-scale. The experimental and numerical results are compared and contrasted to examine the effects of liquid metal feeding in this casting.

Jan 1, 2014

By M. Mirnezami

Pentlandite samples were obtained as Ni concentrate from three Canadian plants and one specimen sample from Ward?s Scientific Establishment. The samples were upgraded using gravity and magnetic separation techniques. Nickel concentration was determined by atomic absorption spectroscopy. Microprobe analysis identified a range in Co and Cu lattice substitutions. All samples exhibited similar zeta potential-pH trend with isoelectric points from 2 to 3, and similar responses to copper activation and addition of xanthate. The similar trends indicate that the variations in lattice substitution identified have no significant effect on electrokinetic properties; that is, samples from any source likely yield results that can be generalized.

Sep 1, 2012

By J. Davitt McAteer

Improvements in the control of coal mine dust have occurred over the past 25 years. Simultaneously, dramatic improvements in coal mine production technology have occurred, particularly at mines where longwall systems are in use. Production rates of longwall systems have quadrupled since 1978, and dust generation on these systems has significantly increased. However, technology to address control of the resulting increased levels of respirable dust has not kept pace, and many high-production longwall sections continue to experience problems with respirable dust control. This paper seeks to enlist the entire mining community in the effort to resolve the health problems presented by excessive levels of respirable coal mine dust, particularly on longwall systems. These systems are expected to increase in number and length, and to generate increasingly higher levels of dust as production rates of these systems continue to rise.

Jan 1, 1995

Agenda ? Where?s Victor ? What work was done to get approval to spend $982m ? a brief history ? Some project facts ? The triple bottom-line accomplishments to date ? Construction A Brief History of Victor ? 1960?s ? first sampling in area ? 1980?s ? stream samples and Victor kimberlite discovered ? 1990?s ? test / bulk samples ? 1999 / 2000 ? large bulk samples ? 2000 / 2003 ? technical evaluation through to Feasibility Study, permitting process

May 1, 2006

By Frederick Beckett

WIDESPREAD experimentation has been conducted in recent years to devise a means of preventing intergranular corrosion in austenitic chromium-nickel steels of the 18 per cent chromium-8 per cent nickel type. Additions of titanium to these steels have been used commercially to overcome this difficulty, as advocated by Krupp of Germany, and the latest literature reference concerning improved details of this method is the article by Bain, Aborn and Rutherford.1 Extensive work by the present authors has shown columbium also to be a particularly valuable agent for inhibiting intergranular corrosion. This paper does not deal with steels that are largely or wholly aus-tenitic ; it is limited to a discussion of the results obtained through the use of titanium and columbium in plain high-chromium steels. For many years titanium has been used commercially by some steelmakers as a cleanser, and to a lesser extent as a hardener, but then in greater per-centages. The results of the work to be described disclose that titanium and columbium in proper proportions act as effective softeners in the high-chromium steels. It is well known that chromium intensifies the hardening effect of carbon in steel; it imparts pronounced air-hardening properties within a rather wide range of chromium content; it provides increased ductility in steels of given strength and hardness; and it enhances resistance to corrosion and oxidation at elevated temperatures to a marked degree. However, in applications requiring a combination of some of these valu-able properties great hardness is frequently a detriment rather than a merit. When the objective is a soft and ductile steel having optimum corrosion resistance, the producer strives for a carbon content as low as commercially practicable.

Jan 1, 1933

[United Engineering Trustees, Inc. W. D. B. MOTTER, JR., '40 A. L. QUENEAU, '41 ALBERT ROBERTS, '41 The Engineering Foundation GEORGE D. BARRON, '44 F. F. COLCORD, '42 A. L. QUENEAU, '41 Engineering Societies Library Board J. W. LAIST, '40 A. B. PARSONS, Ex-officio T. T. READ, '42 F. F. SHARPLESS, '41]

Jan 1, 1940

President Vernal Mortensen, on behalf of 'The RMCMl officers and Executive Board, invites you and your family to attend The 79th Regular Meeting of The Rocky Mountain Coal Mining Institute to be held at Lionshead Village, Vail, Colorado JUNE 26 - JUNE 28, 1983

Jan 1, 1982

By Sidney Katell

Philippine laterite and serpentine ores can be utilized in an electric furnace operation to produce both a steel ingot and a ferronickel product. Figure 1 is a schematic interpretation of a plant in which laterite and serpentine ores are processed separately to obtain both ferronickel and steel as products. A relationship of estimated capital requirements and probable returns has been established in this study for a plant of this type. No attempt has been made to evaluate the risks involved in commercial operations. At such time that the project might be constructed and placed in operation, changes in the costs of material, labor, and power may materially affect the profitability aspects of this study. One of the uncertainties involved in the economic evaluation is the cost of mining, handling, and shipping the ores. It is on this basis that the cost of ore delivered at the plant is assumed to be variable at $3, $5, and $7 per net ton (dry basis). For all other costs a plant site at Iligan City, Philip-pine Islands, has been assumed in preparing the economic characteristics for a plant that will process 2,000 tons4/ of ore per day. Results of the study, for the 2,000-ton plant, with all monetary values expressed in terms of United States currency (Philippine peso valued at $0.50), are summarized as follows:

Jan 1, 1962

By Herbert A. Everest

The idea of economical production is usually associated with large operations, tonnages, and mines, with even larger capital behind them. , Nevertheless many small mines operate in the shadow of large competitors and make a good showing on the capital invested despite larger overhead expenses. For the purpose of this discussion I will divide the cost of production into classes, and specify opposite each the approximate percentagc expended thereon: 1. Labor, including miners and company men; 60 to 75 per cent. 2. Development, including all necessary .yardage, room turning, crosscuts, etc.; 7 to 12 per cent. 3. Deadwork, covering payment to miners for handling falls, draw-slate, or faults, rock work, water, and in general, all nonproductive labor; 3 to 7 per cent. 4. Supplies: mine timbers, oil, brattice material, lumber, cement, and repairs to equipment; 2 to 6 per cent. 5. Expense: management, selling, office, taxes, etc.; 1 to 5 per cent. 6. Depreciation of coal reserves and royalties; 6 to 10 per cent. 7. Depreciation of equipment and interest on capital invested; 1 to 3 per cent. 8. Fuel; ½ to 1½ per cent. The labor cost for a small mine is relatively much lower than for a large mine, particularly the sum paid to company men. Idle-pay expense is cut to a minimum; the labor item being larger than all the others together, a saving in labor makes a decided showing on the total cost. The development charges in small and large mines are about the same. There should be a small showing on the deadwork item in favor of the small mine. The unit cost for supplies is usually small in the small mine. The operator of the small mine finds, owing to a limited tonnage, that the expenses of management, selling, etc., are abnormally high. The item, depreciation of coal reserves and royalties, is usually about

Jan 1, 1916

By H. Ouellet, J. Paraszczak

The pneumatic jackleg drill remains a widely used tool in development and production drilling in many small-scale and narrow-vein underground mining operations in North America. Several drawbacks of this technology (unsatisfactory drilling rate, noise, vibration, presence of oil mist, etc.) and the successful introduction of water-powered jackleg drills in South Africa have stimulated the interest of many mine operators. The performance of both drill types have been tested in a mine site in Northwestern Quebec, Canada. Results of the testing of penetration and advance rates as well as environmental aspects are presented. They show a superiority of water-powered drills, even though the difference in performance is less significant than that reported from South Africa. Further research avenues are proposed to better evaluate the potential of water drilling technology in North American metal mines.

Jan 1, 1995