Search Documents

By G. Paul Anderson

Personnel at Utah Copper Division's Bingham Canyon Mine use current computing technology to compile pertinent production statistics and mining plans. Within hours after the day's work has been completed, concise reports summarizing the previous day's production statistics, actual production vs. budget statistics, the location of shovels relevant to the mining plan, the performance of equipment, equipment next scheduled for maintenance, etc. are available at the terminals located throughout the mine. Also, engineering personnel us e computerized systems, directly from their offices, to assist them in the timely development of mining plans. Timely and concise reporting and planning capabilities such as these assist mine personnel in their efforts to review mining progress and equipment performance, and as such they are in a better position to detect adverse trends early and to seek appropriate resources or make changes in plans or practices.

Jan 1, 1977

By Farlow C. Davis

Magma Copper Company, San Manuel Division, is a wholly owned subsidiary of the Newmont Mining Corporation and located at San Manuel, Arizona. The ore body is a dessiminated porphry copper deposit in quartz monzonite. The primary copper mineral is chalcopyrite with minor amounts of chalcocite, bornite, and covellite. The molybdenite occurs mainly on fracture faces or else associated with quartz veinlets. Pyrite is the major sulphide mineral in gangue. The feed to the concentrator from the underground mine is approximately 0.720 percent copper and 0.029 percent molybdenite. The concentrator has a design capacity of 62,000 tons of ore daily. After a standard ore treatment circuit consisting of grinding, rougher flotation, regrinding, and cleaning, a bulk concentrate is produced which contains 27 to 29 percent copper and 1.0 to 1.2 percent molybdenite. This concentrate is pumped to the molybdenite plant feed thickener.

Jan 1, 1977

By J. W. Francis

The paper about to be presented deals with the relatively straightforward problem of mining two seams of coal separated by an interval. This may seem to be a simple problem but within a twenty-mile radius of Consolidation Coal Company's Burning Star # 4 Mine at least seven (7) different techniques of two-seam mining have either been tried or are now in use. This great variety of techniques should point out that there is no single simple solution; a conclusion often reached when dealing with mining problems. As a background for the techniques presently in use at the Burning Star # 4 Mine, we will briefly outline the other five (5) stripping techniques used in the area: 1. Captain Hine 6360 Range Diagram (Figure 1) 2. Leahy Mine Range Diagram (Figure 2) 3. Burning Star # 3 Mine 5860 Range Diagram (Figure 3) 4. Burning Star # 3 and # 4 Mines Dragline in Pit Range Diagram (Figure 4) 5. Captain Hine 2570/5561 Range Diagram (Figure 5) Each of these techniques was chosen because it provided the best and least expensive alternative for mining the coal under the particular restraints encountered. The Mine Engineers at our Burning Star # 4 Mine, Mr. Dennis Mahan and Mr. William Weaver, will now detail two additional mining techniques, out-lining our experience to date.

Jan 1, 1977

By M. C. Manderson

Phosphate rock is a naturally occurring mineral which, in upgraded form, is important to both agriculture and industry. Phosphate rock normally must be upgraded before it can be assimilated and used. Most of this upgrading is now carried out by chemical means, largely through the reaction of a strong inorganic acid (primarily sulfuric acid) with the phosphate rock to remove major impurities and to produce soluble phosphate. A less important but yet still commercially significant conversion method involves the thermal reduction of phosphate rock at high temperature in an electric furnace to produce elemental phosphorus. The elemental phosphorus is then converted to usable end products. This conversion route is a more costly method than the chemical route, but it has the advantage of being able to produce chemically pure phosphorus derivatives. On the other hand, the use of phosphate rock for agricultural application-primarily for use as a plant nutrient ingredient for fertilizer -does not require the same level of purity. As a consequence, the chemical method of phosphate rock upgrading is the predominant technique for producing phosphatic fertilizers.

Jan 1, 1977

By Robert B. Scott

The Dents Run Demonstration Project near Morgantown, West Virginia, was a cooperative venture involving the U.S. Environmental Protection Agency (EPA), the Water Resources Division (WRD)(1) of the West Virginia Department of Natural Resources (DNR), and the Christopher Division of the Consolidation Coal Company (CONSOL). The object of this project was to demonstrate pollution control methods and to clean up a 37.8 sq km (14.6 sq mi) watershed severely polluted by coal mine drainage. A study entitled, "Water Infiltration Control to Achieve Mine Water Pollution Control" (EPA-R-2-73-142),(2) was conducted by the Cyrus Wm. Rico Division of NUS Corporation for the grantee to determine the feasibility of utilizing remedial techniques to improve the Dents Run water¬shed. This study showed that the Dents Run watershed was discharging as much as 6,870 kg/day (15,100 lb/day) of acid; 1620 kg/day (3,560 lb/day) of iron 17,250 kg/day (38,000 lb/day) of sulfates; and 29,500 kg/day (64,900 lb/day) of total dissolved solids into the Monongahela River near Morgantown, West Virginia. This pollution was caused by the combination of borehole pump discharges from active underground mines, runoff from abandoned strip mines, and runoff from refuse dumps.

Jan 1, 1977

By Ralph H. Light

The recent surge in demand for uranium in the United States, accompanied by a significant increase in price, has resulted in exploration, mining, and milling of lower grade ores than have been exploited in the past. To sustain production levels of uranium with lower grades of feed requires larger tonnage mill operations. Also, because of recent environmental regulations, mill operations must be prepared to properly contain large tonnages of tailings, and recycle, treat and discharge, or evaporate mill effluents. Thus, the advent of lower grade ores, larger tonnage mills, and environmental considerations has made the cost of milling a critical factor in today's uranium production economics in the United States. Three basic milling methods are employed for extracting uranium from ores: 1. Acid leach, countercurrent decantation, and solvent extraction. 2. Acid leach, resin-in-pulp, and Eluex. 3. Alkaline leach, solid-liquor separation, and precipitation. Vanadium often occurs as a valuable by-product in ore deposits located in the Uravan mineral belt in Western Colorado. Also, lesser amounts of vanadium may occur in ores from South Dakota, New Mexico, and Utah. Vanadium is usually extracted in the acid leach solution with uranium and recovered by solvent extraction after recovery of the uranium. By-product recovery of uranium from phosphoric acid and porphyry copper dump leach solutions is now receiving considerable attention. Pilot plant studies are under way and construction of commercial plants is in progress.

Jan 1, 1977

By Ray H. Hario

Background of Erie Mining Company Erie Mining Company, managed by Pickands Mather & Co., is one of the largest producers of iron ore pellets in the United States. The plant is located at Hoyt Lakes, Minnesota, sixty-five miles north of Duluth on the eastern end of the Mesabi Range. Taconite Harbor, situated on the north shore of Lake Superior, is the location of its Power Plant and shipping facilities. (Figure 1) Even though the plant commenced operation in September, 1957, its true be-ginning can be traced back to 1931 when Pickands Mather, in conjunction with Bethlehem Steel Corporation and Youngstown Sheet and Tube Company, began investi-gating the possibilities of processing low grade ores. In 1940, Erie Mining , Company was formed. During the early and middle 1940's, laboratory and pilot plant research was proven to a point where a preliminary plant was deemed feasible and built in 1948. This plant had an annual capacity of 203,200 t (200,000 long tons) per year and, until mid-1952, was the only taconite plant in existence using commercial-size equipment. During this period, ownership of Erie Mining Company was expanded to include Interlake, Inc. and The Steel Company of Canada, Limited. In 1953, authorization was obtained from the owners to build a 7.62 M t (7.5 M long tons) per year plant at a cost in excess of $300,000,000.

Jan 1, 1977

By James A. Clum

The shift to lower grade iron ores, e.g., magnetic taconites, resulted in a shift towards agglomeration practices which could effectively deal with the fine particle sizes generated in the liberation and concentration processes. If the iron bearing particles are not extremely fine, e.g., -1/2 in + 100 mesh, sintering is an adequate means of agglomeration.(1) For the more dispersed ores which result in fine concentrates (more than 80%,-325 mesh), the fine particles must be agglomerated by mechanical means before final hardening, e.g., pelletizing.(1) In the pelletizing process, the essentially mechanical step of balling is extremely critical since the strength of the pre-fired pellet determines the firing conditions. The ability of the proper amount of moisture to hold together a ball of the very fine ore concentrate particles has long been recognized. It is a charac¬teristic of the reaction at the interface between the ionic solid of the ore concentrate and the polar water phase which overlays the solids. The strength in such moisture laden balls, the green strength, is a strong function of the water content (Figure 1) and thus deteriorates rapidly for either a small excess or a small deficiency of moisture. For purposes of commercial practice, the control of green ball handling properties by accurate control of ball moisture is too demanding. Moreover, as the ball is heated towards its hardening tempera¬ture, e.g., 1300'C, there is a substantial drying reaction which would lead to such a low dry strength of the ball that it would not support its own weight much less the weight of other balls in a firing furnace.

Jan 1, 1976

By S. A. Allison, N. P. Finkelstein

After a brief consideration of the properties of zinc sulfide minerals, their flotation characteristics, and their reactions with thiol collectors, the literature on the activation of the minerals by cupric ion, their deactivation by cyanide, and their depression is reviewed in detail. The distinction between deactivation and depression is stressed. The contradictions in the published results and theories are discussed and attempts are made to resolve them. It is concluded that, whereas all aspects of the subject require further investigation, the study of depression is most likely to yield benefits in practice.

Jan 1, 1976

By David H. Dahlem

Uranium is a very mobile element geochemically and is widely dispersed in the lithosphere and hydrosphere. Under oxidizing conditions, it is particularly widespread and appears in all parts of the surface hydrologic system. As a result, this system is a leading candidate in uranium-resource assessments as a new base from which total resources might be estimated. No other sampling system provides the large volume geochemical capture as the continental hydrosphere and is as pervasive and non-selective regardless of rock type. However, the water system as a uranium metal sampler is not without drawbacks as the development of a National sampling program based on it has shown, In spring 1973, the National Uranium Resource Evaluation (NURE) was announced by the U. S. Atomic Energy Commission (AEC). The program was conceived to provide an evaluation of the domestic uranium resources and to identify areas favorable for uranium exploration. The specific NURE objectives are as follows (Bowyer, 1974): 1. To provide a preliminary evaluation of domestic resources and favorable exploration areas by early 1976, based on existing data. 2. To conduct a comprehensive evaluation of domestic resources and favorable areas by 1980, based on advanced state-of-the-art information. 3. To conduct an R&D program into new and improved procedures, equipment, and technology for uranium exploration and assessment. 4. To disseminate information on new exploration concepts and technology.

Jan 1, 1976

By Michael W. Roper

McDermitt Mine is located in northern Humboldt County, Nevada, approximately 5 miles south of the Orgeon-Nevada line (Figure I), within the Opalite Mining District. The operation is a joint venture between Placer Amex Inc., San Francisco, and Mineral Exploration Company Ltd., a limited partnership. The McDermitt orebody possesses a discontinuous, gently dipping blanket-like form within 150 feet of the surface over an area 2200 by 2500 feet. The ore zone averages 20 feet in thickness. The ore minerals consist of cinnabar (HgS) and corderoite (Hg3S2C12) deposited parallel to bedding within argillized, Miocene age, tuffaceous lacustrine sediments. Exploration indicates reserves of approximately 3,000,000 tons ore at a grade of 10 pounds mercury per ton (0.5%) An open pit mining operation utilizing scrapers and bulldozers supplies a concentrator with 700 tons of ore per day. The milling operation consists of autogenous grinding, flotation concentration, concentrate dewatering, and tailing disposal, with a design capacity of 100 tons per hour. The furnace section consists of a concentrate filter, vertical six hearth furnace, mercury condensing system, and exhaust gas scrubbing systems. It has a design capacity of 1000 pounds of dry concentrate per hour. The McDermitt orebody is situated about 2000 feet north of the Cordero Mine which, with adjacent properties, produced over 115,000 flasks (76 pounds per flask) mercury from 1941 to 1972' Two other major mercury producers in the district have been the Bretz and Opalite Mines, located in southern Malheur County, Oregon, within 5 miles of the Oregon-Nevada line. Bretz and Opalite have produced about 14,800 and 12,370 flasks mercury respectively, from 1926 to 1965.1/

Jan 1, 1976

By Hisashi Takahasi, Hironobu Yamazaki

INTRODUCTION The slurry shield method was developed in Japan in 1963 and its experimental tunnelling work was performed in 1965. Although the history of this method is short compared with the conventional shield method in Japan, many companies such as Tekken Construction, Okumura Gumi, Nishimatsu Kensetsu, etc. are engaged at present in developing their own unique system of the slurry shield method. The ground of Japan is generally pervaded with ground water and a large number of districts are covered with soft ground, which is composed of fine sand, silt and clay, etc. Therefore, the conventional shield method has always used compressed air with various kinds of auxiliary methods. However, with this method there have been experienced numerous troubles such as ground settlements, lowering of ground water, oxygen shortage within the tunnel, and so forth. The slurry shield method has been developed with the above background. In recent years remarkable progress has been made from the engineering viewpoint in applying this method to driving various kinds of tunnels such as road tunnels, subways, water supply and sewer tunnels, and other tunnels. Apart from Japan, the slurry shield process using bentonite is carried out in England and Germany. OUTLINE OF SLURRY SHIELD METHOD (1) System of Slurry Shield Method Since the conventional shield method drives tunnels using com-

Jan 1, 1976

By Terry Ramsey, N. Finkelstein

INTRODUCTION The Cyprus Pima Mine is located 20 miles southwest of Tucson, Arizona and about fifty miles north of the Mexican border. The Cyprus Pima orebody was discovered in 1950. Feasibility studies were begun in 1952 employing underground development. Economic studies showed open pit operations to be feasible and stripping of overburden was begun in late 1955. The first ore was reached in October of 1956. A 3,000 tpd (tons per day) concentrator was completed and began operation January 1, 1957. From these beginnings, Cyprus Pima Mining Company has expanded into a plant processing approximately 54,000 to 56,000 tpd. Four expansions were required to bring the plant up to its present capacity. Each expansion incorporated what had been learned in previous operation of the plant. The first expansion was completed in August of 1963 doubling design capacity to 6,000 tpd. In reality, because of softer than average ore, capacity was closer to 8,000 tpd. Expansion I1 was completed in July 1966 bringing design capacity to 18,000 tpd. Expansion I11 followed immediately. Completed in the summer of 1967, this expansion increased the design capacity to 30,000 tpd. The actual dry tonnage of this portion of the operation in 1974 averaged 38,400 tpd, again, primarily as the result of treating softer than average ore. The latest expansion (Expansion IV), completed in December 1971 and utilizing a semi-autogenous grinding circuit, is located in a separate building (Mill 11) approximately 200 yards southeast of the original mill (Mill I). Expansion IV was designed for 14,000 tpd, but after circuit modifications, averaged 16,500 tpd in 1974.

Jan 1, 1976

By E. R. Palowitch

On February 5, 1973, Eastern Associated Coal Corp began shortwall mining in their Federal No. 1 Mine, thus marking the first attempt in the United States at modern shortwall mining, i.e., using a continuous miner to extract coal open-ended from pillars 100 to 200 feet wide, under the protection of a self¬-advancing roof support system (1)1/. Since then, four additional shortwall faces have been operated. While the number of face-years of shortwall experience in the United States is limited, this experience has been anything but uniform. Differences in operator motivation, coalbed parameters, geologic conditions, panel dimensions, and face equipment all have contributed to a wealth of knowledge which can provide guidance for the design of future shortwalls. The purpose of this paper is to review the history of shortwall mining in the United States and to assess its value to the U.S. coal industry in terms of the health and safety of the miner, the percentage of coal recovered, coal production, worker productivity, and relative cost performance.

Jan 1, 1976

By Thomas P. Bligh

Underground and earth covered construction offers a real and significant opportunity for saving energy, reducing building costs, and preserving valuable land resources. These methods lend themselves to mass production techniques without the adverse effects of endless visible repetition and can effect an energy savings of 75% or more.

Jan 1, 1976

By Rodolfo V. de la Cruz

The broad front concept is based on the principle that gasification can be achieved more efficiently and completely from a plane broad combustion front in contrast to other geometries obtainable with current methods of gasification. By restricting combustion to a plane broad front, the ideal condition for gasification is achieved. A plane broad combustion front can be developed in coal by initiating and propagating combustion from the walls of an oriented channel and collecting product gases from another similarly oriented channel. Flow will then be directed along the highest permeability of the coal bed, the available cross-sectional area for fluid flow is substantially increased and, more importantly, the increased flow will be uniformly distributed throughout the whole mass of coal. The success of the method is greatly dependent on how well an initial and rigorous permeability survey and associated corrective measures yield a well-characterized coal bed. To maintain uniform reaction rates, the advance of the combustion front is located with pressure- or mass flow-transducers. These sensors are then connected to servo-controlled mufflers for controlling segments of the fire front which are advancing ahead of others. Finally, the development of the concept to overcome problems of roof collapse, reactant and product gas leakages, and water invasion are briefly summarized.

Jan 1, 1976

By J. F. Cornelsen

Wheel loaders are currently being used in many mining applications for a variety of reasons. While they sometimes provide the lowest loading cost, a primary advantage of the wheel loader versus other loading systems is the increased versatility and mobility inherent in a wheeled vehicle. The advent of large wheel loaders with much greater digging capability enhanced wheel loader usage in severe loading applications. To match these severe conditions, machines offered today were designed to work in these rocky conditions. Soon after the introduction of large wheel loaders into rocky conditions, however, it became apparent to many users that tire costs could at times account for up to 50% of total hourly operating costs on large wheel loaders. In some applications, tire costs alone pushed wheel loader economics beyond the economic feasibility for these jobs, prohibiting their use. In an effort to combat high tire costs, several attempts have been made to put fast moving steel around the tires to provide a strong interface between vehicle and ground. Some earlier attempts involved wrapping tracks around tires. Mechanical difficulties with these systems limited their customer acceptance. Subsequently, the development of protective tire chains came into favor. While tire chains offered added protection to the tires, the cost and maintenance of the chains remains a problem in 1972, the Dystred arrangement was announced for usage on a 6 cu yd wheel loader from Caterpillar Tractor Co. (See Figure 1). The Dystred offered the advantages of greatly increased tread life over conventional tires, as well as resistance to tread puncture in most situations. The arrangement basically employed a track system similar to that used on track type loaders. Pin and bushing wear often caused high hourly operating costs and excessive machine downtime. In addition, the connection of shoes through the use of links, pins, and bushings forced a limitation on too speed of the vehicle.

Jan 1, 1976

By P. Somasundaran, H. S. Hanna

The separation of minerals such as apatite and scheelite by flotation from other salt-type minerals such as calcite is extremely complex owing to the close similarity between their physicochemical properties. In this paper, properties relevant to flotation such as crystal structure, solubility, surface charge and adsorption behavior of important salt-type minerals are examined along with the role of inorganic modifying agents such as silicates and polyphosphates and organic modifying agents such as starch and tannin. Flotation characteristics of these minerals are then reviewed and selected separation problems involving salt-type minerals are discussed.

Jan 1, 1976

By James L. Kelley

The Twin Buttes Mine, one of the more recently developed large copper mines in the United States, is in the Pima Mining District about 25 miles south of Tucson (Figure 1). Other copper producing mines in the district are the Esperanza-Sierrita mining operations, five miles to the southwest and the Pima and Mission Mines, seven miles to the north. Copper metal production from the Pima District in 1974 totaled 276,728 tons and the Twin Buttes Mine's share was 46,473 tons or about 17%.' The district's copper production was 31% of Arizona State's production and nearly 17% of the production from the United states in 1974. The Pima District thus ranks as one of the major copper producing areas in this country.

Jan 1, 1976

By J. Herbst, C. C. Harris, G. Poling

The modern flotation machine is the result of decades of adaptation to a changing environment. In response to falling ore grades, and rising demands, tonnages, and costs, attempts are now underway to expand machine capacity by about an order of magnitude, while at the same time maintaining metallurgical performance and if possible reducing power requirements per unit throughput. Empiricism and rule-of-thumb have guided flotation machine design and scale-up, and a current need is for the development of a fundamental understanding of the subject. Design features of current and some obsolete machines are discussed in an attempt to outline those principles of flotation machine design which are known, and to delineate topics needing investigation. Attempts are made to distinguish between essential machine features required by the physics of the process, those features resulting from speculative theories on the flotation mechanisms, and features which are essentially styling variations. Topics discussed include: flotation mechanisms; tank and impeller/ stator design and dimensions; impeller speed; aeration; power consumption; particle suspension and dispersion; capacity; correlation of design parameters and operating variables; scale-up and scale- down; materials flow; problems in machine design. Design details and operating data for several current machine families are provided in Appendices.

Jan 1, 1976