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By M. J. Shulman

The exploration geologist is frequently confronted with the problem of allocating his resources, i.e., exploration effort, among various areas. Although geologic, financial and logistic considerations influence the allocations, no objective criterion is usually given which determines the optimal allocation. Maximizing the gain in information, subject to the constraint of a limited exploration budget, is used as the objective criterion to allocate reconnaissance effort. The information function, I(x), quantifies the expected gain in information, where x is the scale of exploration. To calculate the information function the following data are necessary: 1. The probability that an area contains a target must be estimated. 2. For each scale of exploration, four probabilities must be estimated: the probability that an anomaly is produced by the target; the probability that an anomaly is not produced by the target; the probability that an anomaly is produced by a barren area (false signal); and the probability that an anomaly is not produced by a barren area. Alternatively, the barren area can be replaced by other targets. To illustrate the allocation model, the optimal number of line kilometers/line spacing per area is determined for an aeromagnetic survey for iron in Egypt.

Jan 1, 1985

By Gonzalo Barrientos

Instrumentation and observation of real mine structures helped define the pressure arch at White Pine, a large copper mine in sedimentary rocks. The pressure arch, which is attributed to inherent horizontal stresses, is used in the design of mine openings.

Jan 1, 1969

By Lester F. Sanford

Located in the Santa Cruz Mountains eight miles north of the City of Santa Cruz and inland from the coast, lies a typical beach sand consisting of quartz and feldspar grains of the Santa Harguerita formation of the middle tertiary age, Reserves of over twenty-five million tons of this sand form the deposit presently being harvested by Pacific Cement & Aggregates, a Division of Lone Star Cement Corporation. The deposit ranges from 350 to 700 feet above sea level and varies from two percent passing a 200 mesh screen to ten percent retained on a ¼ inch screen (see Figure '#l) In 1929 the first sand plant went into operation at this site. It was replaced by a new plant in 1937. This plant continued operation until the middle of 1965 when replaced by the modern plant described in this paper.

Jan 1, 1967

By M. V. K. Reddy

A differential flotation process for recovery of molybdenite from chalcopyrite concentrates produced at Rakha Copper Concentrator, Bihar, India has been developed. The flowsheet is simple, and relatively inexpensive and yields a concentrate assaying not less than 45% Mo at a recovery of minimum 65%, starting from chalcopyrite concentrate assaying 0.4% Mo. The consumption of the main depressant used, sodium sulphide is not more than 25 kg/t of copper concentrate processed without steaming; and with steaming the consumption was brought dorm to less than 12 kg/t of concentrate. Further reduction is possible by optimization of steaming parameters. Based on laboratory experimental results, further tests were conducted on concentrator site and process design parameters were established for design of commercial plant.

Jan 1, 1982

By D. H. Y. Tang

3-dimensional finite element modeling was performed to analyze the safety factors of stump pillars, 18 ft. x 18 ft., under a cover of 208 ft. using the Modified Drucker-Prager Theory. The results showed that they are safe. This was substantiated by underground observations. But the safety factors predicted by three commonly used pillar design formulae showed otherwise. This paper discusses the modeling details and anlyze the difference between 3-D FEM method and traditional pillar design formulae.

Jan 1, 1987

By David A. Summers

Historically, the presence of methane has been a problem, mainly in and around the working areas of active coal mines, and only in these area has drainage been considered. Drainage, where practical, has been achieved through the drilling of holes forward into the coal and the surrounding strata from the working area. These holes generally have been short in length, although where methane drain- age operations around a longwall face have been undertaken, the holes have had to be longer in order to adequately drain from the center of the face into the access gate roads. In recent years, attempts have been made to degasify the coal seams in advance of mining, without disruption of the mining cycle. This is done by drilling much longer horizontal holes through the coal in advance of the working area. Under the aegis of the Federal Government, methods have also been developed for draining seams of coal of their methane content in advance of mining, but from shafts sunk from the surface, without using the active area of the mine as the location for the drill holes.

Jan 1, 1979

By Floyd L. Prestridge

A test of a solvent extraction (SX) pilot plant featuring electrostatic coalescence has been performed by C-E Natco at the United Nuclear uranium mill in Church Rock, New Mexico. Operating in parallel with the United Nuclear SX system, the C-E Natco electrostatic SX pilot plant exhibited comparable extraction efficiency to the United Nuclear system which utilizes conventional mixer-settlers. The C-E Natco pilot, however, operated at much higher liquid loading rates and exhibited less communication between the aqueous phases of the extraction and stripping circuits.

Jan 1, 1983

By James R. Yancik

Economical and environmental pressures necessitates the recovery of the minus 0.6 mm x 0 coal generated by todays mining machines. This size fraction, dependent on the type of production unit, maintenance of the production unit, and the Hardgrove Grindability Index, can be a substantial portion of the Preparation Plant feed. Historically, dewatering of the collected 0.6 mm x 0 product has often been rather expensive and marginal with respect to surface moisture reduction. The Freeman United Crown II Preparation Plant more than adequately demonstrates that the continuous scroll type centrifuge can dewater 0.6 mm x 0 coal to acceptable surface moisture levels.

Jan 1, 1981

By R. Kingman

At Tahawus, New York, National Lead Company operates the MacIntyre Development, This operation produces 10,600 long tons of ore per day for crushing, 8,000 tons for grinding and processing, of which 2,000 tons per day reports as tailing for disposal in an impounded tailing storage area. With the change in location of mining operations at the mine of National Lead Company, a change in the area for tailing disposal by gravity became necessary, This called for the diversion of tailing from near the proposed mining area to areas more favorable for long term storage by gravity. In order to avail the operation of maximum storage for the future, a uniform minimum grade launder was proposed with a minimum velocity and maximum outfall at the discharge. Disposal of the gravity and flotation tailing was formerly accomplished in open ditch from the concentrator at a grade of 1.5 percent. It was desirable to design the new launder at grades less than this for the greater storage elevation.

Jan 1, 1966

By R. E. Coleman

Infrared spectroscopy, employing the internal reflection method in conjunction with solvent extraction techniques, was used to study surface reactions between aqueous potassium ethyl xanthate and both sphalerite and precipitated zinc sulfide each activated with aqueous copper sulfate. It was determined that reactions occurring on the sphalerite and zinc sulfide surfaces produce the reaction products ethyl dixanthogen and cuprous ethyl xanthate, some of which are ad¬sorbed on the mineral surfaces. According to infrared spectra showing the reaction to form these compounds, the ratio of cuprous ethyl xanthate to ethyl dixanthogen adsorbed is directly related to pH of the flotation solution.

Jan 1, 1967

By S. C. Gambrell

The practice of roof bolting in world-wide mining activities is very well known, having been used extensively for over 50 years. The theory of roof bolting, however, is extremely complicated, and only recently have investigators attempted to explain the theory. Discussions of the mechanics involved usually originate with actual experience gained in the use of roof bolts. The use of roof bolting in mines has grown dramatically during the period after World War II. Other scientific techniques have emerged in mining research in the same period. Included in these techniques is the study of rock mechanics, which has now developed into an impressive research effort. Various means of analysis have been utilized in rock mechanics, most of which reduce rock to an isotropic, homogeneous medium that could be easily analyzed. Because of the inherent nature of rock, direct research methods frequently are not certain to provide satisfactory results, whereas indirect methods must idealize a given set of conditions in order to be useful. One of the indirect research methods is the photoelastic technique, the theory of which is widely known and needs no elaboration. This technique has been used extensively to examine the state of stress around underground openings of various shapes in certain types of strata, and under various configurations (1,2,3,4,5). Much of the theory used now in the design of underground openings is based on photoelastic evaluation. Unless rigorously applied, how- ever, the photoelastic technique is at best a qualitative means of evaluation. It is useful, however, for unusual examples which can be more or less easily analyzed to obtain the overall state of stress which could not be determined by other means. Little work has been done in the photoelastic analysis of rock bolting, mainly because direct research techniques have proven useful.

Jan 1, 1970

By S. L. Bessinger

In the design and specification of a ventilation system for an underground mine, it is necessary to make reasonably accurate estimates of the pressure losses in the various airways of the mine. These estimates can be made with little difficulty for open airways with simple geometric cross-sections, such as those cut by continuous miners or tunnel-boring machines. The situation is much different on a longwall face, where the airway's complex geometric cross-section and the presence in the airway of obstructing equipment having a variety of shapes make it difficult, if not impossible, to estimate pressure loss using traditional methods of calculation. Headlosses in mine entries are calculated using Atkinson's Equation (1). [H = (English) H = (SI) (1)] where H = pressure ,loss, in. of H20 (Pa) K = friction factor, lbf min2/ft"4(Kg/m3) P = perimeter, ft. (m) L = airway length, ft (m) Q = airflow quantity, ft3/min (m3/sec) A = flow cross sectional area, ft2 (m2)

Jan 1, 1987

By Don Muraoka

Soda ash, anhydrous sodium carbonate, is produced from underground trona deposits occurring in the Green River Sasin of Southwestern Wyoming. Stauffer Chemical Company of Wyoming, a jointly owned subsidiary of Stauffer Chemical Company and Rocky Mountain Energy Company mines the trona and processes it into high-purity soda ash. Stauffer Wyoming has been in operation since 1962 with a current production capacity of approximately 2.2 million tonnes (2 million short tons) per year of soda ash. The Refinery has five parallel operating trains utilizing the monohydrate process which is a series of chemical processing unit operations described below. Screening and Crushing the Trona The mined trona is brought to the surface in two counterbalancing ore skips. The ore is initially screened with Refinery feed, less than peagravel size material, entering the process while the oversize is conveyed to an outdoor stockpile.

Jan 1, 1984

By Frank Littlewood

Research over a period of ten years resulted in the development of an efficient automatic ore feed regulator for mill grinding circuits. The facilities that a large mining company gave the inventor made it possible to develop the control apparatus under actual mill operating conditions. The Littlewood Regulator automatically determines the optimum power input at which a mill should operate, and controls the rate of ore feed to maintain this optimum power for any predetermined operating conditions. It is to be noted that the optimum power varies with many factors such as ball, rod, and mill liner wear, therefore the regulator is designed to automaticall keep itself properly calibrated. To efficiently control the rate of ore feed of a mill grinding circuit the regulator analyzes and interprets all power changes of the mill motor. The critical electrical measurements for this type of control are made possible by a patented device that has the ability to detect and integrate minute changes in power. In ball mill operation the small but significant changes in power serve to select the rate of ore feed in accordance with the rate at which ore accumulation takes place within the mill, and in rod mill operation they serve to select the rate of ore feed in accordance with the degree of rod spread.

Jan 1, 1970

By L. F. McHugh

Molybdenum conversion operations discussed in this paper encompass all major operations commercially used to upgrade molybdenite concentrates to various refined products. More emphasis is placed on discussion of key pyrometallurgical unit operations. These operations conducted at Metec include roasting, sublimation, and hydrogen reduction

Jan 1, 1986

By W. H. Reck

he need in the mineral industry to treat increasingly larger tonnages has caused a demand for larger installations. In order to limit the size of these installations, it was necessary to build machines with larger capacities or production output. With this aim in mind, WEMCO Division developed the No. 120 WEMCO-Fagergren Flotation Cell in 1967. These large cells have been compared to conventional smaller cells from the standpoint of metallurgical performance, investment costs, power requirements and maintenance requirements in many plants throughout the world. Figure 1 illustrates the size difference of the No. 120 when compared with the No. 66. All results to date indicate the large No. 120 cell surpasses the smaller machines in all categories.

Jan 1, 1969

By Robert C. Meaders

In this paper, autogenous grinding is defined as the combined crushing and grinder of a run-of-mine ore or the product from a primary crusher, in a single stage grinding mill. This is in contrast to pebble milling where pebbles are substituted for steel balls to grind a ball mill size feed. The autogenous mill to compete in efficiency with normal methods of crushing and grinding must not use greater power, must use less labor and steel, and must produce an equivalent product.

Jan 1, 1963

By L. B. Sand

Zeolite ores and protores occur in extensive deposits in the western United States. A recent paper1 describes the general geology and mineralogy of these deposits and their geographical distribution. The zeolites are alkali-and silica-rich varieties of mordenite, erionite, chabazite, phillipsite, ferrierite, and clinoptilolite. The zeolites were formed primarily by the alteration of Tertiary pyroclastics and in some cases by the alteration of Tertiary siliceous flows. The altered pyroclastics are of principal economic significance. The volcanic ash and tuff were deposited in restricted basins where saline lakes developed. The pyroclastic layers often are of uniform thickness and macrotexture for miles from the center of the basin and interfinger with detritals toward the edge of the basin. Thus, in addition to unaltered volcanic glass, the gangue minerals in these deposits include some combination of primary minerals in the pyroclastics, alteration products of the glass other than the zeolite of interest, and detrital minerals. The gangue minerals include quartz, cristobalite, tridymite, opal, orthoclase, micro cline, montmorillonite, hornblende, calcite, gypsum, thenardite, and iron oxides. The presence of one or two other zeolites in the layer in some cases also is considered as gangue. There are no quantitative mineralogical descriptions of these deposits in the literature; this: study was undertaken to obtain this information. As the zeolites occur as single crystals, crystal aggregates, zeolite-glass particles and in cemented aggregates, a method of study was chosen which would give the size distribution of the zeolite in these types of association. For this reason dispersion of the rock was achieved with a minimum of grinding action. This excluded those zeolite ores, silicified by hydrothermal altera¬tion which require grinding. A relative index of the ease of disaggregation of the zeolite ores thus is obtained, but the technique does not give absolute size distributions of the zeolite crystals. Samples were selected for analysis which do not necessarily represent the best ore horizon, but do represent a spectrum of zeolite occurrences from significant deposits. They are mordenite, erionite, and phillipsite from Rome, Oregon, clinoptilolite and chabazite from Reese River, Nevada, chabazite (and var. herschelite) from Bowie, Arizona, erionite from Jersey Valley, Nevada, and erionite from Pine Valley, Navada.

Jan 1, 1967

By L. F. Moran

Rod Mill Grinding is an important and fundamental step of the concentrating process in Marcona Mining Company's plants at San Nicolas Bay, Peru. This Paper summarizes experience in regard to consumption and specifications of grinding rods based on twelve years operation of the grinding circuits. Marcona operates a Plant complex producing 10 million tons per year of beneficiated iron ore products including pellets, sinter feed concentrates, high grade concentrates as slurry and filter cake. Processing capacity is 15 million tons per crude ore per year of which 13 million are around in rod mills. A continuous testing program was carried out over the years. Initial studies of daily production operations gave indications of poor grinding rates in relation to mill size and power consumption, with a high rate of broken liners and rods. The broken nieces of rod were not automatically expelled and further interfered with grinding efficiency, while replacement of liners caused unwarranted do n tire. Objectives were to improve all phases of the Process by studying the effects of changes in rod diameter, length, alloy constituents, and volume and weight of rod charge.

Jan 1, 1975

By Ole Bernt Lile

The well-known method of C. Schlumberger, Mise a la masse, has become a popular method again in ore exploration in Norway after many years of dominance by EM methods. The principle of the method is to place a current electrode within a conducting ore body and remove the other electrode to a great distance. (Figure 1). Often it is convenient to place a return electrode on each side of the area explored. The method is best suited for ores that have a conductivity contrast, relative to the host rock, of several orders of magnitude. The potential drop within the ore itself may then be neglected and the ore contact against the rock may be regarded as an equipotential surface. One may say that the ore body is electrically charged. Therefore, in Norway the method is now most often called the Charged Potential Method (abbreviated CP). The most usual way to present data obtained by CP, is to draw equipotential lines based on potential values measured in profiles on the surface. Such a potential map gives an immediate picture of the outcrop of the ore and also indicates disturbing conductors in the neighborhood. It may also be useful to look at the potential profile and its horizontal gradient along the strike axis of the potential map. These profiles may give information about the strike length of the conductor. Figure 2 shows the theoretical profiles of potential and gradient along the axis of the potential picture on the surface over a charged line plunging 15 degrees. It may be seen that the maximum of the gradient curve clearly indicates the lower end of the conductor. Figure 3 shows the corresponding pro- files over a charged line plunging 45 degrees. In this case, it is not possible to see an indication on the curve of the lower end. Only curve fitting could help interpretation in this case.

Jan 1, 1975