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By R. W. Snyder, H. J. Ramey

This paper presents the results of applying the Buckley-Leverett' displacement theory to petroleum reservoirs consisting of a finite number of layers. The layers are assumed to communicate only in the wellbores, and the reservoir may be represented as a linear system. Most previous investigations of this nature were limited by assumptions and by inconsistent calculation techniques. This study improves on previous work by applying the Buckley-Lev-erett displacement theory to a noncommunicating layered system where permeability, porosity, initial saturation, residual saturation and relative permeability vary from layer to layer in a logical and consistent manner. Gravity and capillary-pressure effects are neglected. A modification of the Higgins-Leighton calculation method was used in this study. Waterflood predictions were made with all properties varying, and then with only permeability varying using several inability ratios. These results were compared with the Stiles and Dykstra-Parsons predictions. It is shown that the latter methods generally give poor values for the breakthrough recovery and pessimistic predictions for the performance after breakthrough. Similar results were obtained for a gas-displacement case. lNTRODUCTION Field experience with immiscible displacement usually shows constant producing conditions until breakthrough of the displacing fluid. Then oil production continues at increasing displacing-to-displaced fluid ratios until the economic limit is reached. Three different ideal mechanisms are known that will produce this behavior: (1) relative permeability effects as described by Buckley-Leverett frontal advance theory,' (2) vertical stratification as considered by Stiles,2 Dykstra and Parsons5 and others and (3) different path lengths involved in areal (two-dimensional) flow between wells as described by Dyes et al.4 Without question, a combination of these factors modified by formation heterogeneity and other known and unknown factors actually does control the behavior of real systems. This paper presents results of an investigation of certain factors that should affect performance but which have received little attention to date. In 1944, Law5 demonstrated that porosity and perme- ability are often found to have normal and logarithmic-normal distributions, respectively. throughout cored intervals in natural formations. This led to the concept of the noncommunicating, multilayered reservoir model for immiscible displacement. This model assumes that the reservoir is composed of a number of layers that communicate only at the wellbores. Each layer is individually homogeneous, but may be different from every other layer. Stiles' presented one of the earliest applications of this model to waterflood performance. In addition, Stiles assumed that the initial saturations and relative permeabilities were the same for each layer, porosity was the same. displacement was piston-like, fluids were incompressible and injection into each layer was proportional to that layer's permeability capacity (permeability-thickness product). The last assumption would be true if the mobility ratio for the displacement were unity.21 Dykstra and Parsons" used the same model as Stiles, but rigorously included mobility ratios other than unity for piston-like displacement. Dykstra and Parsons used their general result to produce charts for log-normal permeability distributions between layers. Similarly, Muskat6 Pub1ished analytical solutions for linear and exponential permeability distributions. In 1959, Roberts' described a scheme for calculating water-drive performance for the noncommunicating, layered reservoir model which considered two-phase flow in the displaced region. Roberts used the same model and assumed that the injection rate into a layer was proportional to that layer's permeability capacity, but that flood front locations could be evaluated from the Dykstra-Parsons results. These assumptions are inconsistent, and a material balance cannot be maintained except for a mobility ratio of unity. At the same time, Kufus and Lynch8 coupled Buckley-Leverett displacement theory with the layered model to provide an improvement of the Dykstra-Parsons method that was consistent. In 1960, Higgins and Leighton9 resented a numerical method for calculating waterflood performance also considering two-phase flow in the displaced region. The result was used to investigate variation in absolute permeability and oil viscosity. An excellent, detailed history of using the noncomrnunicating, layered reservoir model was presented by Nielsen.'" The preceding techniques (and many related ones) were similar in that differences in initial saturations, residual saturations and relative permeabilities from layer to layer were neglected. It is well known that the irreducible water saturation is an important function of absolute permeability. Calhoun11 showed that the irreducible water saturation

By Donald W. Jenkinson

Many types of gravity concentration equipment have been employed throughout the years since beneficiation of materials has been practiced. These include jigs, concentrating tables, vanners, sluice boxes, heavy media cyclones, and static heavy media separators or sink-float separators. In the beneficiation of coarse aggregate for concrete mixes, static heavy media separation employs the same principles of gravity concentration involved in separating lead-zinc ore, dolomite, limestone, diamonds, manganese, iron, and many other minerals. In aggregate beneficiation, as in most metallurgical operations, sink or heavy fraction is the product. The undesirable portion of coarse gravel is usually much lighter than the desirable particles, allowing most aggregates to be amenable to gravity concentration methods.

Jan 7, 1960

By L. J. W. JONES, B. H. Bennetts

THE removal and disposition of large quantities of slag from blast-furnaces is a question of great importance in the design of works, and various methods have been devised, from time to time, in order to take the best advantage of local conditions. In blast-furnaces treating copper-ores or lead-ores, it is necessary to use a fore-hearth, or matte-settler, in order to collect, if possible, every particle of valuable matte which, otherwise, would be carried away with the out-flowing slag. In all of these cases, the slag overflowing from the fore-hearth is caught in pots and then conveyed to the dump. If the slag is of no value and the necessary water-supply, grade and dumping-space be available, the disposal of the slag by granulation is by far the cheapest and best method. It involves a minimum of labor, both at the furnace and at the dump; and it permits a continuous outflow of slag from the furnace- or settling-pot, thereby avoiding frequent tapping and plugging. When the conditions of water-supply, grade and dump-area will not permit slag-granulation, there are three other methods in common use. Large steel or cast-iron pots, or ladles, supported by trunnions on heavy cars which are moved by steam-, air-, or electric locomotives; some type of mechanical conveyor; and small pots or buggies moved by hand. The pots on the slag-cars are generally circular or oval in shape, and each one contains from 15 to 40 or more en. ft. of molten material. The mechanism for tipping them is more or less elaborate; and, as a general rule, the slag is dumped alongside, and quite close to, the track. They are more or less costly to construct and to keep in repair, and, in order to use them to best advantage, it is necessary to have a good elevation of track above the dumping-ground; furthermore, the rails must be of heavy weight, and well-supported on heavy ties. Other disadvantages of the slag-pots are-1, that the slag cannot be removed continuously

Mar 1, 1905

By G. F. Ziesing, E. F. Loros, T. J. Smolik

The Golden Sunlight mine began full-scale operations in February 1983 with a conventional open-pit mine supplying 4536 t/d (5000 stpd) of milling ore. After crushing and grinding, the ore is leached with cyanide solution in large agitation tanks. Soluble gold values are removed from the sand portion of the leached ore with five stages of countercurrent washing cyclones. After thickening of the cyclone-wash overflow slurry, the gold is recovered in activated carbon with a six-stage carbon-in-pulp circuit and a five-stage upflow carbon column circuit. Pressure stripping, carbon reactivation, electrowinning, and refining complete the gold recovery process.

Jan 1, 1985

By John E. Caffrey

The Federal Coal Mine Health and Safety Act of 1969, Title III, Interim Mandatory Safety Standards for Underground Coal Mines, is devoted to protection of workmen from hazards of electricity, roof, ribs, fire, and explosions. The Rules and Regulations, promulgated through the Federal Register, provide the criteria by which the law would be administered. Certain aspects of the law involve features that could be summarized under entrapment. These are sections 303 (y)(1), (y)(2) and 317 (f) (4). This paper is intended through the use of the figures to describe how these features of the law affect mine ventilation design, what are the problems, and possible solutions to them.

Jan 1, 1974

By Charles C. Upham

Not long ago a few acres of coal-land in the Connellsville region of Pennsylvania were sold at the rate of $1500 per acre. While this was doubtless a " fancy " price, affected by some consideration other than that of the simple value of the coal in the tract. it may fairly be taken as an indication of a progressive and already perceptible diminution in the economically available coal-resources of that famous region. This piesent and prospective development emphasizes the cominercial importance of producing, from the inferior coals of other districts, coke suitable for use in the blast-furnace. Many of the bituminous coals of the United States would meet this requirement if sufficiently freed, before coking, from slate and sulphur, and, so far as the removal of slate is concerned, the problem is not specially difficult. There are many forms of trough-, jig- and gravity-washers which easily and cheaply reduce the slate from as much as 20 per cent. in the coal to as little as 8 per cent. in the coke. But to reduce the sulphur from 4 per cent. in the coal (a proportion sometimes encountered) to 1 per cent. or less in the coke (the proportion demanded by the furnace-manager) is not so easy. It may be safely assumed as familiarly known to members of the Institute interested in this subject that, of the three forms in which sulphur usually exists in coal (hydrogen sulphide, calcium sulphate and pyrite), the pyrite is the one to be removed, RS far as possible, by mechanical treatment before coking. Hydrogen sulphide, being volatile, is expelled in

Jan 1, 1899

By Georges Fleming

The new burden preparation plant of ARBED Luxembourg, Belval- Works, which comprises 2 large sinter strands, is equipped with a process computer connected to a comprehensive process control system. The main jobs done by the computer are the control of the raw mix composition, and thus the monitoring of the stacking of the piles at the bed-blending plant, and the optimization of the sintering process. Among the optimization criteria, quality is considered as the most important. The sinter quality, characterized by the mechanical strength and the state of oxidation, is controlled automatically. The control system which we have dev loped is highly reliable; as a result, the standard deviation of Fe++, for the control of bivalent iron, has been narrowed to 0,5 % and the standard deviation of the IS0 + 6, 3 mm index, for the control of the mechanical strength, has been kept within 1 %. Thanks to these results, the combined savings in fuel consumption at the sinter plant and at the blast furnaces have covered the total costs of the optimization project in 2 years. By applying automatic control, the optimum driving range of the sinter plant has been defined and these new guidelines are used for monitoring the sinter production.

Jan 1, 1977

By G. S. Rice

Undoubtedly there is a serious problem through the juxtapostion of gas and oil wells and coal mines, not only at the present time, but possibly of far more serious import for the future.

Jan 1, 1915

By George Dyott

THERE are many who may consider a paper on aerial transport and its possibilities in connection with mining operations somewhat premature. Nevertheless, as there is considerable interest manifested in the subject at the present time and the amount of information available is both scanty and misleading, the result of investigations in some of the remote mining districts of Peru will not only be of interest, but will help crystallize, in the minds of many, a more correct conception of just what an airplane is capable of doing. By way of introduction, I will deal briefly with the types and varieties of airplanes now in use, and the developments of the past few years, pointing out -how the war has encouraged machines. of startling performance rather than those of commercial utility. I will then take up aerial transport in general, showing how airplanes may be employed to advantage and the particular kind of work for which they are best suited. This will cover a preliminary discussion on cost, which is one of the most important phases of the subject. In passing on from a general survey of the situation, I will deal with actual observations and investigations made in Peru. This part of the paper will be divided into two sections: the first, covering the mountainous districts of the Andes (Tamboras, Pataz, Soledad, etc.); the second, the alluvial deposits in the rivers, Chinchipe, Santiago, and Napo. In each case the question of atmospheric conditions, altitude, landing grounds, motors, cost, etc. will be discussed. While it is my intention to deal chiefly with mining in the sections mentioned, the same line of argument will be applicable to any other districts in the world. In conclusion, I will touch on the possibilities of using lighter-than-air craft of the Zeppelin type, followed by a summary of the most salient points brought out in the paper.

Jan 2, 1919

By Carl E. Williams

A computer program has been developed which will locate the economic limits of an open-pit mine and also designate which ore and waste blocks to remove per period (years or quarters) to maximize the discounted cash flow over the life of the mine. A set of operating constraints was carefully selected so that the following functions are controlled: (1) the removal of overburden by preproduction and production stripping exposes a fixed amount of ore; (2) a constant manpower and equipment usage is insured (except for trucks); (3) the tons of mill ore, concentrate, and stockpiled ore are regulated; (4) the number and extent of shovel moves are minimized; (5) the grindability of the rock is considered as well as the variable haulage cost caused by variable haulage distances and depths; and (6) the size of the shovel working area is regulated. As many as 50 alternate schedules can be produced per computer run. They are arranged in descending order starting with the best solution.

Jan 1, 1975

By T. P. Meloy

Economic constraints require that the optimum mineral processing circuit be chosen for a given ore and then the circuit be optimized. Meloy (1983) developed a general methodology for finding the best circuit for a given ore. This analysis, however, did not specify a way to optimize the best circuit once it has been found. To optimize a given circuit, it is necessary to find the optimum set points for each unit operation in the circuit. Developed in this paper is a simple, general equation which, when applied, defines the optimum separation set point for the maximum mineral grade. This general equation, Equation (4), applies not only to simple ores, but also to complex ores containing variable and locked particles. Currently, the optimum set point can be calculated only for heavy media and cyclone circuits, because they are the only unit operations for which the necessary data has been measured.

Jan 1, 1984

By Charles J. Speake, Dennis R. Haley, Thomas E. Finch

The tremendous interest in development of surface- mineable coal throughout the western US will bring with it a requirement for many manhours spent in mine planning and equipment selection. One problem in particular that engineers face during the planning stage is that of selecting a dragline with the proper reach capability. Reach, expressed in terms of operating radius and effective spoil radius, can be a major problem when mining deeper than originally planned, and aside from production requirements, the ability of a particular dragline to spoil across a strip is essential. Determination of dragline reach needed for a multiple-seam western coal operation can be simplified by use of an equation derived below that eliminates the need for tedious and repetitive calculations.

Jan 5, 1977

By R. E. Caffrey, S. K. Tung

This paper reports some of tile results obtained from the vapor-phase reaction of a volatile metal halide with a hydrogen and carbon dioxide mixture in an epitaxial-deposition chamber. Tile oxides deposited were silicon di oxide, aluminum oxide, mix-12tves of silicon di oxide and aluminum oxide, and titanium oxide. Tile deposited silicon di oxide was transparent and amorphous and had a density and index of refraction similar to that of thermally grown silica. The vote of etching was determined. The thickness uniformily of the oxide was found to he related to the aerodynamics of the system employed for deposition. A clear silicon di oxide layer was found to be dependent upon the quality of the surface prior- to aside deposition. The process parameters affecting the vale 0f deposition were studied. The deposited aluminum oxide was transparent, holy-cl-ysttrlliile, and possessed remarkable resistance to hydrofluoric acids. By measurements it has been shown that the deposited aluminum oxide is both insulating and continuous. ThE cleanness of an epitaxial silicon surface both immediately after growth and during device processing is thought to be one of the major factors contributing to the electrical properties of a final device. The advantage of protecting an epitaxial silicon wafer surface with an oxide grown immediately following epitaxial deposition is obvious. The successive growth of silicon and silicon diox- ide in epitaxial apparatus was studied by Steinmaier and loem.' They formed silicon dioxide films on a silicon substrate by adding water vapor to the silicon tetrachloride-hydrogen mixture in an epitaxial system. They also showed that, when a gaseous mixture of carbon dioxide and hydrogen is injected, water is formed only at elevated temperatures. Thus the reaction of water vapor and silicon tetra-: chloride will be limited to a relatively small reaction zone near the substrate. This paper reports some results of oxide films obtained from a vapor-phase reaction of a volatile metal halide with a hydrogen and carbon dioxide mixture in an epitaxial-deposition chamber. EXPERIMENTAL APPARATUS The system employed for the oxide deposition is a modification of a vertical system used to produce high-resistivity silicon layers by the hydrogen reduction of silicon tetrachloride.' The modification shown schematically in Fig. 1 consisted of an additional carbon dioxide gas stream which entered the reaction chamber with a hydrogen stream saturated with metal halide vapor and a diluting hydrogen stream. Silicon dioxide layers made from a mixture of silicon tetrachloride, hydrogen, and carbon dioxide gases will be discussed first. EXPERIMENTAL RESULTS 1) Factors Affecting the Rate of Oxidation. An epitaxial silicon layer was deposited by the hydrogen reduction of silicon tetrachloride on an n-type,

Jan 1, 1965

By Leonard Larson

Before direct or wet smelting of copper concentrate was adopted at the McGill smelter, in November 1932, actual furnace smelting tests had indicated the possibility of smelting between 400 and 500 dry tons of wet concentrate per furnace day. Under calcine smelting practice, solid charge up to 116; tons per furnace day had been smelted. It was realized that in order to smelt wet tonnage in equivalent amount per furnace day, a higher combustion rate and higher temperatures would be necessary within the furnace. The reverberatory furnace considered in the following history of wet-smelting developments is a unit rz7 ft. long and 28 ft. 8 in. wide, inside dimensions (Fig. I). This furnace was in operation through all the years from 1934 to date. During 1933, the first complete year of direct or wet smelting, the plant was on a curtailed copper-production basis and the available concentrate tonnage to be smelted was limited. The average coal-firing rate in the reverberatory for that year was 124 tons per furnace day. This rate was lower than the average for years immediately preceding under calcine smelting practice. With this coal-firing rate, an average of 391 tons of new metal-bearing material (N.M.B.M.) was smelted per furnace day. The following year, 1934, the coal-firing rate was increased to 139 tons and the smelting rate increased to 448 tons N.M.B.M. per fur- nace day. In succeeding years production required that either more tonnage be smelted in the one furnace or an additional furnace be placed in operation. With this situation confronting the smelting department, it was determined to continue the program of increasing the coal-firing rate and also to intensify the temperature in the smelting zone of the furnace. In order to accoinplish this, various improvements were made to aid combustion within the furnace. The burners were enlarged, increased primary air was supplied to the burners, and changes were made in the furnace to enlarge and streamline the outlet, thus increasing the furnace draft. The coal-firing rate was increased yearly until 1937 when the firing rate averaged 210 tons per furnace day. Thereafter the firing rate fluctuated according to character of charge and production requirements. In general, however, with the same character of charge the smelting rate increased in almost direct proportion to the increases in firing rate. Table I shows the average coal-firing rate, smelting rate, and other pertinent data for the direct-smelting years of 1933 to 1939 inclusive, the first fivc months of 1940 and May 1940. These figures show that the water evaporated in the waste-heat boilers was in almost direct proportion to the coal-firing rate. In making comparisons in the table, it should be borne in mind that the figures, except for May 1940, are yearly averages. Because of changes in character of charge or operating conditions, records for individual days or even months

Jan 1, 1944

By C. S. Barrett

Measurements of impact hardness of ß-brass are reported for a wide range of temperatures. Abrupt softening is observed when temperatures are raised above 425ºC, accompanied by abrupt widening of deformation bands in the deformed samples. Current theories of these effects, and of the lack of twinning, are discussed. HARDNESS1 and flow stress2 of a sample of ß-brass at room temperature depend upon the temperature from which the sample is quenched. It is believed that the state of order in the superlattice is responsible for this dependence—not the order within individual domains, which reaches near-perfection during the quench from any temperature, but the size of the domains which is presumed to be a minimum in the hardest samples, those quenched from temperatures near the critical disordering temperature of 468°C. Measurements of impact hardness made over a range of temperatures and microstructural studies of samples deformed both rapidly and slowly over a range of temperatures, reported herein, should be of aid in reaching a more complete understanding of the underlying factors controlling the properties of the disordered body-centered-cubic material and the ordered material. It appears that such data have not

Jan 1, 1955

By Gerd L. Klinkenborg

Ever see a curvaceous conveyor belt? Reports from Europe indicate appreciable interest in a relatively new system that has revamped some of the concepts of what single-belt conveyors can and cannot do. Invented about ten years ago by two German engineers, the so-called "Serpentix" single-belt conveyor, patented by the firm of Albert Stuebbe, is now operating at more than 100 foreign installations. The system boasts of three unique capabilities-climbing, turning and side discharge of load-which give it, the advantages of a multiple-unit system. The new conveyor needs only one drive station (which may be combined with tensioning), and except for the immediate area of the drive, is capable of conveying loads over a non-linear route. Transfer points are eliminated, and discharge points, which may be located anywhere along the route, may be operated automatically or manually, by direct or remote control.

Jan 1, 1962

By Howard Martens

HARDNESS behavior of commercial titanium alloys following various heat treating processes has been studied for some time. However, the hardness of such alloys following a definite measured cooling rate from the single phase region has not been reported. Therefore, a series of experiments was conducted on the three alloys: 150A, 130A, and 130B. The manufacturers gave the following compositions for these alloys: Ti 150A: 0.045 pct C, 0.069 pctN, 1.32 pct Fe, and 2.68 pct Cr; RC 130A: 0.13 pct C, and 7.9 pct Mn: RC 130B: 0.11 pct C, 3.5 pct Mn, and 3.2 pct Al. A method similar to that described by Greninger was used, and the cooling rate varied from 6" to 9200°F per sec. The cooling curves were recorded on a Speedomax or a recording oscillograph, and the hardness was measured on a Tukon hardness tester using a Knoop indentor with a 500 g load. The curves of Fig. 1 show the variation of the Knoop hardness with the cooling rate for these three alloys. The alloys 150A and 130B behave in a similar manner, as shown by the curves. In neither of these alloys was it; possible to retain the high temperature ß phases, even by the most rapid cooling rate. The slow cooling rates produced a typical ß structure in these two alloys. As the cooling was increased, these phalses became more finely dispersed and the hardness increased. This increase in hardness with cooling rate continued for the 150A alloy until a rate of approximately 1000°F per sec was reached. At this cooling rate, the structure showed the result of the dif-fusionless transformation of the ß solid solution to the supersaturated a solid solution which is referred as a'. At higher cooling rates, the structure showed no change and the hardness showed no great change, as indicated by the dotted portion of the curve in Fig. 1. The increase in hardness with cooling rate for the 130B alloy continued until a rate of approximately 300°F per sec was reached. At this rate the structure was produced and the maximum hardness was reached. Higher cooling rates caused no marked changes in hardness or structure. The alloy 130A behaved in a different manner, as shown by the curve of Fig. 1. The structure produced by the slow cooling rates was a typical a+ß structure, which became finely dispersed as the cooling rate increased. The hardness increased with cooling rate and reached a maximum at approximately 120°F per sec. At this maximum hardness, the structure showed the result of the decomposition of the ß phase into the ß + phase structure. As the cooling rate was increased, smaller amounts of the at phase were formed and the alloy became softer. For cooling rates of 3000°F per sec and above, the high temperature ß phase was retained and the hardness of the alloy did not change markedly. Acknowledgment This paper presents the results of one phase of research carried out at the Jet Propulsion Laboratory, California Institute of Technology, under Contract No. DA-04-495-0rd 18, sponsored by the Dept. of the Army, Ordnance Corps.

Jan 1, 1957

By Samuel G. Lasky

In a recent discussion of the mineral resource position of the United States, published in the September 1946 issue of Mining and Metallurgy, I expressed my conviction that ore deposits of major magnitude remain to be found in this country and that they can be found without undue cost or risk. I stated with equally strong conviction, however, that if the search for such deposits is to be notably successful there must be, among other things, a reorientation of philosophy and mental habits on the part of many who would be connected with the search. I would now like to enlarge upon that statement. For such a purpose. we may consider that the search for concealed deposits involves four steps: (I) the selection of areas for further study; (2) the geologic study of those areas in order to verify their general favorableness and in order to choose specific patches worthy of being tested; (3) the decision whether exploration of the patches chosen should actually be undertaken; and (4) the exploration itself. The first two steps and the last rest upon the judgment and skill of geologists. The third step—the decision whether exploration should be undertaken—rests upon management. In my opinion, the philosophies or principles that have guided these steps in the past are now inadequate. Philosophy of Appraisal If concealed deposits are to be looked for, how are the places of their possible concealment to be recognized? Every known mineralized area still in the prospect stage has already been repeatedly examined and condemned. As for those deposits concealed under later rocks, there may be not even the clue of a near-by outcrop for engineer or geologist to consider. flow then, are we to know where to look? Geologic Probability In the early days of our mining industry, ore deposits were found mainly by "following the ore." Later, as such opportunities decreased, and as geologic concepts became accepted, deposits were found more and more through use of the notion of geologic probability. Many instances can be cited wherein a courageous operator ignored the implications of the unsuccessful efforts of predecessors and, solely by "following the ore," converted a discredited prospect into a good mine. Similar experience doubtless will occur in the future, but "following the ore" will no longer in itself find new deposits in the amount needed. Nor can enough new ore be expected from further use of the notion of geologic probability, for, as mentioned above, our known prospects have already been ex-

Jan 1, 1949

By Samuel G. Lasky

In a recent discussion of the mineral resource position of the United States, published in the September 1946 issue of Mining and Metallurgy, I expressed my conviction that ore deposits of major magnitude remain to be found in this country and that they can be found without undue cost or risk. I stated with equally strong conviction, however, that if the search for such deposits is to be notably successful there must be, among other things, a reorientation of philosophy and mental habits on the part of many who would be connected with the search. I would now like to enlarge upon that statement. For such a purpose. we may consider that the search for concealed deposits involves four steps: (I) the selection of areas for further study; (2) the geologic study of those areas in order to verify their general favorableness and in order to choose specific patches worthy of being tested; (3) the decision whether exploration of the patches chosen should actually be undertaken; and (4) the exploration itself. The first two steps and the last rest upon the judgment and skill of geologists. The third step—the decision whether exploration should be undertaken—rests upon management. In my opinion, the philosophies or principles that have guided these steps in the past are now inadequate. Philosophy of Appraisal If concealed deposits are to be looked for, how are the places of their possible concealment to be recognized? Every known mineralized area still in the prospect stage has already been repeatedly examined and condemned. As for those deposits concealed under later rocks, there may be not even the clue of a near-by outcrop for engineer or geologist to consider. flow then, are we to know where to look? Geologic Probability In the early days of our mining industry, ore deposits were found mainly by "following the ore." Later, as such opportunities decreased, and as geologic concepts became accepted, deposits were found more and more through use of the notion of geologic probability. Many instances can be cited wherein a courageous operator ignored the implications of the unsuccessful efforts of predecessors and, solely by "following the ore," converted a discredited prospect into a good mine. Similar experience doubtless will occur in the future, but "following the ore" will no longer in itself find new deposits in the amount needed. Nor can enough new ore be expected from further use of the notion of geologic probability, for, as mentioned above, our known prospects have already been ex-

Jan 1, 1949

By J. H. Ffolliott

ALL, of the prospects that later became mines in the Ducktown district were discovered prior to 1854, and most of them were exploited before the turn of the century. It is the purpose of this paper to discuss the exploration and current development of the Boyd mine, discovered in 1853. It is in the heart of the Ducktown Basin, in the extreme southeast corner of Tennessee, and is being developed by the Tennessee Copper Co. GEOLOGY The geology of the district has been fully described in many papers. Probably the most complete description is that of Emmons and Laney.1 The rocks of the district are considered to be of Lower Cambrian age and are a part of the Great Smoky formation, consisting mainly of interbedded graywacke and mica schist, in which the ore deposits occur. The ore bodies are included in an area of about 6 sq. miles, 2 miles wide and 3 miles long striking N. 45° E., and may be grouped into three general lode systems. The northeast system, including the East Tennessee, London and Burra Burra ore bodies, and the southeast group, comprising the Polk County, Mary and Calloway deposits, are all characterized by relatively high copper values. The ores of the central system, which includes the Eureka, Boyd, Old Tennessee and Cherokee ore bodies, are low in copper and high in total sulphide content. Boyd mine lies roughly halfway between the Old Tennessee and Eureka ore bodies [ ] and 2500 ft. in the hanging wall of the Burra Burra mine (Fig. 1). Elevation of the surface is about 1725 ft. above sea

Jan 1, 1942