Search Documents

By Robert Keeney

AT the beginning of the use of the electric furnace, for the manufacture of calcium carbide and ferro-alloys, experimental work was conducted in it upon the production of steel from iron ore. Stassano performed two series of experiments in Italy, the first1 during the years 1898 to 1902, and the second2 in 1908. For many years steel and wrought iron have been produced directly from ore on a small scale in the forge, crucible, and reverberatory furnaces; but the difficulties attending the operation of these combustion processes have prevented their adoption to any great extent. The combustion

Jan 2, 1914

By L. A. Wright

Union Carbide Corp.'s Pine Creek mine in California is the largest known tungsten deposit in the West, and the producer of well over five million tons or ore to date. Located 21 miles northwest of Bishop on the eastern slope of the Sierra Nevadas, the mine is also a source of molybdenum and copper and it was the quest for the molybdenum which caused the mine management to undertake a pillar recovery program. GEOLOGY For a full discussion on the geology of the Bishop tungsten district and the Pine Creek mine, reference is made to Special Report 47, Calif. Div. of Mines. The deposit lies along a 3000 ft segment of the contact between quartz monzonite and the marble that flanks the west side of the north portion of the Pine Creek pendant. All the components of a typical contact metamorphic tungsten deposit are developed; these consist of tactite, light colored calc-silicate rock and quartzose rock.

Jan 5, 1964

By Charles Holland

MINERS have observed for many years that as pillars are removed in mining operations the pillar adjacent to those mined out frequently shoes evidence of being compressed. Although this has been known for a long time, little or no information is available concerning how far from the mining area the pillars are affected in this manner, or how great is the magnitude of the compression, or concerning the manner or char-acteristic way in which the compression acts, as the mined-out area approaches the pillar under consideration. The writer made a study of pillar action in the Mona mine of the Arkwright Coal Co., partly to get some quantitative data on the deformation of coal-mine pillars. The study was not completed but it seems best to present the information that was obtained.

Jan 1, 1937

By G. B. Pryde

IN laying out a new mine, provision should be made for the ultimate recovery of as much coal in any given bed as is consistent with safety and economic mining. Though each mining district, if not each individual mine, has problems that make the adoption of any hard and fast rule impossible, certain basic principles are adapted to practically all conditions. One of the most vital is the provision for barrier pillars of ample dimensions to protect all haulage and permanent airways during the life of the property. Where conditions permit, room or butt entries should be driven on the strike or slightly to the raise and rooms turned at right angles to said entries. The width of rooms and room pillars is governed by conditions prevailing in the mines under consideration. A safe rule is: In advancing, extract not more than 25 to 35 per cent. of the entire seam; in exceptionally thick seams, considerably less than 25 per cent. should be taken. Rooms usually vary in length from 250 to 350 ft. (76 to 106 m.) with a barrier pillar of not less than 80 ft. between the last crosscut in rooms, and entry and room crosscuts should be driven on sights; room crosscuts should be driven at right angles to the direction of the rooms. In mines under excessively heavy cover or where roof conditions are bad, rooms should be driven narrow and a barrier pillar left at every fifth to tenth room. This barrier should be of such dimensions that one or more rooms, with pillars of usual dimensions, may be turned through it when the barrier is no longer required. Two of the advantages of such a barrier are the protection of the panel against peaking of roof load during pillar extraction and the easy closing of the entry in case of fire.

Jan 2, 1921

By W. E. Hess

At the Vesta mines of Jones & Laughlin Steel Corp. on the Monongahela River, 35 miles south of Pittsburgh, JCM Joy continuous miners and 6-SC shuttle cars are used for pillar extraction in the Pittsburgh seam. The entire output of the mine goes into coke production for blast furnace operation. Because this coal, which has high metallurgical qualities, is close to the place where it is ultimately used and because supplies with these properties are being rapidly depleted, complete recovery is essential wherever possible. With the introduction of the continuous miner into the Vesta mines, therefore, it was necessary to plan a project which would facilitate development and provide for future total recovery. Use of continuous miners under varying conditions has necessitated changes with regard to roof action, method of development, method of pillar attack, and innovations in mining. In the following pages these changes will be discussed.

Jan 2, 1955

By J. C. Irvine

Meramec Mining Co., a joint venture of Bethlehem Steel Corp. and St. Joe Minerals Corp., mines and pelletizes iron ore at the Pea Ridge mine. The Pea Ridge property is located near Sullivan, Miss., about 113 km (70 miles) southwest of St. Louis. The ore body was delineated in the mid-1950's by St. Joe during a lead exploration program. The first test holes drilled on the Pea Ridge magnetic anomaly revealed the presence of a large magnetite deposit; further drilling in 1956 and 1957 confirmed that the ore body was mineable. Meramec Mining Co. was incorporated in 1957 and shaft sinking began late in the year. Production began in April 1964. The ore body is overlain by about 400 m (1300 ft) of flat-bedded sediments. It is tabular, about 800 m (2600 ft) long and up to 180 m (1600 ft) thick. It dips about 80° and is of unknown depth. The ore is mainly high-grade magnetite with small zones of specular hematite. The wall rock is a Precambrian rhyolite porphyry.

Jan 9, 1976

By Arthur H. Wilson

INTRODUCTION The stratified deposits of the coal measures are strongly laminated and exhibit a high degree of anisotropy. The stronger rocks in the sequence possess joints and shear planes, the coal itself has a closely spaced cleat pattern, and the softer rocks are capable of flowage. To calculate the general distribution of stress in such a medium by analytical means based on the theories of elasticity is difficult, and has not yet been satisfactorily achieved. Nevertheless, some estimate is required to aid the design of supporting pillars and permit the correct siting of main roadways. This is important when laying out districts for longwall extraction, particularly if the results of previous experience are not available in the area to be exploited. Although elastic stress analyses, including finite element analysis, may not be very appropriate, the basic equilibrium conditions will apply. These, combined with logical deductions based on observed fact, can be used to give appropriate solutions which experience has shown to be capable of giving acceptable results in the design of pillars associated with longwall workings. Constancy of cover load is one of the fundamental equilibrium conditions. When mining below ground, the cover is not removed, therefore the average vertical load after mining must equal the initial cover load. If mining produces a local decrease in load in any particular area, it must be balanced by an equivalent increase in load nearby. Bending moments in the strata above are capable of compensating for the turning moments which this redistribution of load may produce, and the existence of a "rear abutment", much discussed between 1950 and 1960, is not an essential equilibrium requirement. If the distribution of vertical stress is plotted three-dimensionally, then the load on any small element of area 6S will be o.6S, where o is the average stress. This also represents the volume of the "stress envelope" which lies above 6s. By making the individual areas very small and summing together, it is possible to show that the total load on any particular area can be found by calculating the volume of the stress distribution envelope which lies above it. A similar consideration applies to the stress distribution in a two-dimensional situation. If the stress distribution curve is drawn, then the area below any portion will represent the load per unit depth of section. For example, Figure 1 represents the possible stress distribution across a caved waste and associated ribside. If AW is the load deficiency [ ] These loads must balance, therefore area AW = AS, and if one can be deduced, the other will then be known. Attempts have been made to measure stress distributions in the ribside and in the waste. Stress meters in soft rocks such as the coal measures have not proved very effective. Those based on the measurement of strain have experienced difficulty in strata capable of flowage, high modulus stress meters suffer from cross sensitivity, and overcoming to obtain absolute stress is almost impossible. Load measuring instruments left in the waste have also proved largely ineffective, mainly as a result of the difficulties of protecting the leads in a caving environment. Certain logical deductions, however, can be made about the conditions in the waste. CONDITIONS IN THE WASTE Over the area of the waste the surface trough which develops is remarkably consistent, despite the geological variations which occur from coalfield to coalfield. In Britain, an accuracy of 210% is claimed for subsidence prediction based on width of extraction and depth of cover (National Coal Board, 1975). The subsidence in the USA is more variable (Adamek and Jeran, 1981), particularly with respect to the point of inflexion, but the same general pattern occurs. At the shallower depths the somewhat harder roof strata involved in the caving in the USA may suffer less compaction, giving reduced values of subsidence. Subsidence at intermediate horizons below the surface

Jan 1, 1982

By W. R. Wawersik, A. M. Starfield

The classical approach to room-and-pillar mine design in flat-lying, tabular ore bodies limits itself to the special case of an extensive and uniform system of pillars. Each pillar is then assumed to carry the same load, and this load is readily calculated from the extraction ratio e and the weight of the overburden : average pillar stress = [ ] where H is depth below surface, p is average density of overburden, and g is acceleration due to gravity. The average pillar load does not, in this case, depend on the mechanical properties of either the pillars or the surrounding rock mass. The only relevant physical property of the pillars is their compressive strength, (i.e., the maximum load-carrying ability) and the criterion for design is that the average pillar stress should not exceed this strength. The wider the pillar, the greater its strength, and the design of the system is concerned with reconciling this increase in strength to the requirements of roof control so as to optimize extraction.l Any departure from a uniform distribution of pillar loads invalidates the classical approach. Such departures may, for instance, be brought about by: (1) a nonuniform layout of pillars and remnants resulting from selective mining, (2) variations in geology which result in gross changes in the properties of part of the seam or the host rock, (3) deliberate use of abutment support in narrow panels, and (4) the use of yielding pillars for local roof support in conjunction with barrier pillars to carry the overburden load.

Jan 1, 1972

By S. Harries Daddow

THE INSUFFICIENCY OF PILLARS OF COAL FOR THE PURPOSES DESIGNED-THE FRUITFUL CAUSE OF DANGER, EXPENSE, AND WASTE-THE PROOF OF INSECURITY-SUBSTITUTE FOR PILLARS OF COAL-PILLARS AND PANELS COMPARED. PILLARS of coal, which are designed and left in our mines as the means of supporting the overlying strata of our coal-beds, and securing protection to the miner, are not only insufficient for these purposes, but the most objectionable means of providing the desired security to life and property. Though such pillars are left at a great sacrifice to the owners of collieries and mineral lands, loss to the resources of the commonwealth, and ultimate privation to the public, they defeat the very ends for which they are designed, and not only do not secure life and property, but are as fatal to the one as ruinous to the other. Propositions.-It is eminently desirable that the resources of the commonwealth-particularly our mineral wealth in anthracite coal, which is limited-should be carefully utilized and made available, not only to ourselves but our posterity. Yet, speaking as an anthracite miner, we are wasting our resources of coal with a recklessness which will bring ruin in the end, unless checked, not only to the iron manufacturing interests of the East, whose chief protection, in competition with the iron-masters of the interior, lies in a cheap and abundant supply of anthracite furnace fuel, but to the best interests of the State. It is, consequently, not only desirable, but necessary, that we should realize every ton of coal from each colliery, and every acre of coal area, in order that our resources may not be wasted at present, to breed want in the future, and to realize the greatest present benefit to the mine, the miner, and the public. It is not possible, however, to secure these ends, if we continue to leave, as we do at present, from one-third to one-half our resources in anthracite as pillars in our mines. But our sad bills of mortality, our abandoned collieries, and our yawning mountain-sides, present ample evidence that even this enormous waste of coal, though left for the purpose of protecting life and property, is as totally inade-

Jan 1, 1873

By James E. Reynolds, Duane N. Goens, Charles W. Kenney

A process for recovery of lead and silver from high-grade lead sulfide concentrates by dry chlorination and fused salt electrolysis was developed by Hazen Research, Inc., for Cyprus Metallurgical Processes Corporation (Cymet). A pilot plant was designed to handle mixed sulfide concentrates and to produce high purity lead in a 2,000- ampere capacity fused salt cell. Design of the dry chlorinator and fused salt cell is discussed. Integration of modern chlorine caustic cells for reagent generation was evaluated in the pilot plant program. Also included in the paper will be process chemistry; pilot plant flowsheet, equipment and data: by-product and impurity removal: and scaleup and economic considerations for the process.

Jan 1, 1977

By Robert M. Lewis

Silica sand deposits are usually contaminated with various heavy minerals which must be removed to make the silica useful for flat-glass production. Research was undertaken to develop a better procedure for producing flat-glass-grade silica. Batch tests revealed that anionic detergent-type reagents had considerable promise. Pilot plant tests were run to compare the effectiveness of anionic detergent-type reagents with petroleum sulfonate in the removal of heavy minerals from a potential glass sand. Another slightly modified commercial procedure also was evaluated for comparison. In the pilot plant trials, utilization of the anionic detergent potation reagents did not produce quite as good results as petroleum sulfonate. It is likely that flat glass-grade sand could be produced with the anionic detergent reagents if more extensive pilot plant tests were conducted.

Jan 1, 1977

By John G. Reid

Subsequent to extensive laboratory testing, a pilot plant was constructed to further evaluate the process for high temperature copper cementation on iron in continuous operation. The plant was designed to produce 1 to 10 tons of copper a day depending on the copper concentration of the pregnant solution. The energy required to operate the plant was very low as virtually all the solution heating requirements were supplied by the exothermic cementation reaction. The results demonstrated that this, process has potential for commercial application with respect to operating considerations, and yields a copper product in excess of 95% copper at high reaction rates, low iron consumption, and at least 98% copper recovery from solution.

Jan 1, 1976

By M. E. Kukura, L. G. Stevens

As described in a previous presentation and in various patents, the Minerals Sciences Div. of UOP, Inc., has developed a modified reduction roast ammonia leach process for nickel laterites. This process gave considerable improvement in extraction over the conventional Caron approach as practiced in Cuba, Australia, and the Philippines. This greatly improved response was obtained with a wide variety of ores. This paper describes the pilot plant flowsheet and reviews the design features unique to the UOP process. Operating results from pilot plant campaigns are compared with bench scale results. Operating problems and their solutions are also discussed.

Jan 1, 1976

By Mark E. Smith

To allow recovery of gold values from ore which cannot be economically processed by the mill, Pinson Mining Company has developed a run-of-mine heap leaching facility. In order to estimate production scale recovery rates, reagent consumption and percolation problems, a series of pilot scale tests were performed. These tests typically consisted of 3,630 metric tons (mt) (4,000 tons) leach ore at a grade of .94 to 1.56 grams Au per metric ton (g/mt) (.03 to .05 oz. Au/ton). Test heaps were loader stacked to approximately 4.3 meters (14 feet) high. Information from these pilot scale tests was employed in the design and construction of the production scale facility. Parameters such as spray rate, leach cycle, sprinkler placement, and barren solution chemistry were determined by the pilot tests.

Jan 1, 1984

By S. F. Ravitz, S. R. Zimmerely

High-grade cobalt concentrates were produced from the complex Blackbird ore with very good recovery in continuous pilot-plant operations in which a low-grade bulk cobaltite-pyrite flotation concentrate was roasted in a multiple-hearth furnace to oxidize the pyrite selectively. The cobaltite was floated from the resulting calcine.

Jan 10, 1950

By C. K. Stoddard, R. G. Knickerbocker, E. T. Leidigh, R. R. Lloyd, K. L. Mattingly

DURING July 1941, a study was initiated at the Boulder City Experiment Station of the Bureau of Mines on proposed methods for the production of magnesium metal. The major emphasis was placed upon development of a process that required minimum reliance upon the sale or utilization of by-products to meet economic competition from processes already in operation. It was felt that only through the development of such a process could the nation's vast resources of dolomite and low-grade magnesite that are contiguous to large blocks of firm power be properly utilized. During the course of exploratory tests on electrolytic and electrothermic processes, it was noted that fused magnesium chloride would dissolve appreciable quantities of magnesium oxide in the presence of small amounts of boric oxide. The use of such a solution in an electrolytic cell employing a graphite anode and a steel cathode caused the deposition of magnesium at the cathode and the evolution of carbon oxides at the anode with minimal amounts of chlorine The original results were sufficiently encouraging to justify further investigation on an enlarged scale, since the process seemed to present advantages over other proposed processes 1.2 for the direct use of magnesium oxide in a fused chloride electrolyte. When work on a larger scale was initiated it was found that the addition of boric oxide to the electrolyte resulted in dispersion of metal and constantly decreasing current efficiency. However, by this time solubility studies had shown that magnesium oxide was sufficiently soluble in magnesium chloride melts containing no boric oxide to effect the same anode reactions, and development work was continued. Over two years of development work have been carried out, and considerable work remains to be done. It is desired to emphasize the fact that many of the difficulties met require a long period of continuous operation before they become serious enough to affect operation of the cell. ACKNOWLEDGMENTS This paper is one of many reporting on various aspects of the Bureau of Mines' program initiated in August 1939 by passage of the Strategic Minerals Act the scope of which was greatly expanded by subsequent legislation. Some of these papers are published as war minerals reports. others as bulletins, technical papers, reports of investigations

Jan 1, 1945

By John H. Bassarear, Harold W. Sorstokke

The never ending search for lower unit costs led Pima Min- ing Company's management to study various alternatives shortly after Expansion III was completed. A considerable amount of ore as- saying 0.35% copper had to be moved to maintain mill feed at the desired 0.54 % copper content, which is the overall grade of the ore reserves. Detailed investigations revealed that for a period of 8-10 yr, about 14,000 tpd of 0.30- 0.40% copper ore pass the crusher enroute to the low grade stockpile. This ore is very hard, very clean and has to be mined at a uniform rate. A six month pilot plant run testing autogenous and semi-autogenous grinding gave encouraging results. Since no mining costs need to be charged to the extra tonnage (it had to be mined anyway), the major costs would be against milling. The decision to proceed was made in August 1970. Fluor-Utah was retained as the prime design construction contractor. Ore was put through the new mill on December 29, 1971. By March of 1972 design tonnage

Jan 5, 1973

By John H. Bassarear, Harold W. Sorstokke

The fourth expansion within 10 years was completed at Pima Mining Co. during late 1971. The first three expansions increased capacity from 3000 to over 40,000 tpd. Conventional crushing and grinding practices were used in the original 10-section mill. Total capacity is now approximately 54,000 tpd. Expansion IV includes a new primary crusher and a radial stacker conveyor that feeds stockpiles for both concentrators. The -8-in. primary product is fed from the reclaim stockpile to the new mill by two pairs of pan feeders, each pair feeding a separate conveyor. The reclaim conveyors each feed 28 X 12-ft semiautogenous mills operating with an 8% charge of 4-in. balls. Autogenous discharge is screened, oversize recycling by conveyors, with -%-in. undersize feeding two conventional 161/2 X 19-ft ball mills in closed circuit with cyclones. Cyclone overflow is pumped to three parallel rows of 300-400 cu ft flotation cells. Rougher and scavenger concentrates produced are sent to the original mill for regrind, cleaning, and filtering.

Jan 1, 1974

By George A. Komadina

The Pima mine lies 20 miles southwest of Tucson, Ariz., and 45 miles north of the Mexican border. It is located on a hillside overlooking Tucson, which allows for gravity flow generally through the plant. In April 1965, an expansion was authorized by the mine's board of directors. This will triple the mine's output to 18,000 tpd. At this rate, it is estimated that there is sufficient ore to last 20 years.

Jan 6, 1965

By Michael D. Martin

Pima Mining Co., managed and 50% owned by Cyprus Mines Corp., with Union Oil of California and Utah Construction and Mining Co. sharing equally in the remaining 50%, is a medium-sized open pit copper mine situated in the very active Twin Buttes mining district, 22 miles southwest of Tucson, Ariz. In 1965, a three-part paper described Pima's growth from 3,000 tpd (tons per day) in 1957 to 18,000 tpd nine years later. Within months of the completion of the last stage of this sixfold expansion, studies were being wound up on a further major expansion, this time to 30,000 tpd. This expansion is now a reality. Details of the enlargement of the crushing and concentrating plant have already been covered in a recent paper, and the present paper will cover pit engineering studies, equipment selection, and current operating practices. Key factors in top management's decision to go ahead with the latest expansion were: (1) a substantial increase in ore reserves coupled with an appreciable decrease in grade, (2) a generally favorable outlook for copper prices, and (3) better than expected performance by mine and mill equipment purchased for the expansion to 18,000 tpd. For the pit design studies, particular attention was paid to pit stability, a problem area at times in the past, and to the location and grades of the final ramp systems. In drawing up the long-term mining schedule, the attitude of the ore, with the higher grade material lying deeper, was a critical factor and considerable stockpiling of lower-grade material in the early years had to be accepted in order to keep the annual grades close to the average of the orebody. The expansion called for major purchases of mine equipment in order to gear up the mine for delivery of 30,000 tpd to the concentrator. In general, equipment chosen was of types already proven at Pima, al- though there were some exceptions to this practice. One of these excep-

Jan 1, 1969