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By Ed L. Reed

INTRODUCTION The most efficient development of a mine plan for in-situ leach (solution) mining must be based upon an understanding and application of the basic hydrologic characteristics of the aquifer containing the mineral values. Similar knowledge of the overlying and underlying aquifers is also required. This paper summarizes the hydrologic properties of a ground water system and describes some of the more important conclusions that can be drawn from a detailed hydrologic study. Basic to an in-situ leach mining operation is the flow of fluids through porous rock from place to place. The degree to which these movements of fluids can be predicted and controlled determines to a large extent the efficiency of the ore recovery operation. HYDROLOGIC CHARACTERISTICS The principal hydrologic characteristics of a porous medium (sand, intercrystalline or vuggy limestone, fractured basalt, etc.) influencing fluid flow are: Porosity Permeability Thickness Transmissivity Storage Coefficient Water Table or Piezometric Surface Hydraulic Gradient Briefly these parameters can be defined as follows: Porosity: the void space in a unit volume of rock. Permeability: the ease with which fluid will flow through a porous medium. Thickness: the height of the formation which contributes flow to a well. Transmissivity the product of permeability and thickness. Storage Coefficient. the volume of water released from the reservoir per unit drop in head. Water Table or Piezometric Surface: the level at which water will stand in a well: the top of the zone of saturation in a gravity aquifer; the height to which water will rise above the confining bed in an artesian system. Hydraulic Gradient the direction and rate of dip of the water table or piezometric surface. A properly designed pumping test will provide numbers for transmissivity, storage coefficient and the water table or piezometric surface. Drilling records or bore hole logs provide aquifer thickness data from which permeability can be calculated. Ground elevations applied to measured depths to water enable one to calculate the direction and rate of dip of the water table or piezometric surface. Porosity can be determined from core analyses or density logs; it can be estimated from grain size distribution curves in sands or sandstones and from, microscopic examination of drill cuttings derived from carbonate aquifers. CHARACTERISTICS OF FLUID FLOW The characteristics of fluid flow that are vital to an understanding of leach mining include: Velocity Direction Flow Path Radius of Influence The velocity of fluid flow is of paramount importance in planning and managing a leach mine operation. Velocity is directly proportional to the permeability and hydraulic gradient and indirectly proportional to the effective porosity The permeability determined from a pumping test is an average value for the interval stressed; therefore, the velocity calculated from data derived from a pumping test represents average conditions. Under actual conditions discrete portions of the aquifer open to the screen may have permeabilities several times as high as the average and minimum permeabilities only a fraction as much. An injection front will be irregular to the extent that the reservoir rocks are heterogeneous An important consideration is the direct relationship of velocity to head. A reservoir which would provide a velocity of 0.05 feet per day under conditions of equilibrium would allow fluid flow of 36 feet per day under typical leaching operations, i.e. 12-foot injection head, 12-foot pumping level and 50 foot five-spot pattern. The direction of fluid flow is also vital both inside and outside the mine area. In the pattern area the direction of flow is determined by the geometry of the well field and can be approximated by mathematical modeling techniques. Outside the well field, the direction of flow is influenced by both natural conditions and the degree to which production and injection volumes are balanced. Flow directions (and velocities) are critical to an understanding of leachate excursions and their control.

Jan 1, 1979

By T. M. Chance

THE necessity of adopting improved methods of bituminous-coal cleaning at. many collieries is recognized by all familiar with the requirements of the various bituminous markets, the approaching exhaustion of the relatively small areas of inherently high-grade-coal, and the difficulties encountered in maintaining a uniform grade of shipped coal. Many coals lose much of their .market value by degradation in size, and, in any event, if the cleaning process is to contemplate wetting the coal the minimum amount of fines should be produced to prevent sludge losses and the difficulties resulting from the moisture content of the wet fines. Hence, it is desirable to by-pass the finer coal dry, if this can be done without too-great increase in impurities in the final product. This procedure has given excellent results at washeries at Steelton and in the Birmingham district. If the raw fines are too impure to permit admixture with the washed coal, it may be desirable to use dry methods, such as the pneumatic table, rather than to wet an excessive amount of such fine material. The bituminous-coal cleaning plant should therefore: (1) Treat the raw coal in as large lump as possible and without sizing; (2) eliminate degradation of pure coal -in the washing process; (3) avoid wetting the fines, if possible; (4) produce washed coal practically free from mechanically included impurities; (5) crush the refuse and retreat, if much intergrown coal, and refuse are present, rather than precrush the entire feed. Ordinary gravimetric methods of concentration, whether hydraulic or pneumatic, will not meet all of these requirements, as the large lump includes those sizes. now sent to the picking tables, if it is desired to treat material of such size. I do not know of any prior commercial method, based on the falling velocity, coefficient of friction or resistance to fracture, that permits the production of clean coal varying in size from, say, 10-in. lump to 1/16 -in. mesh.

Jan 6, 1924

By Frank Doxey

AT Dominion Steel & Coal Corp. it has long been recognized that continued mechanization of mine operations is necessary in the Pictou, Cumberland, and Sidney coal fields of Nova Scotia. The varied physical conditions in these fields call for special consideration of individual cases before planning is finalized. Because standard equipment cannot be procured which would operate successfully, many experiments have been necessary over the years to keep pace with the progress made in other countries. There are two mines, producing 2000 tpd, located in the Pictou coal field. The field is badly distorted and crossed by many faults. Seams are highly inclined and irregular and vary in thickness from 5 to 40 ft. Entries are difficult to maintain because of squeezing of the coal. ribs and movement of the roof and pavement. Output from the three operating mines in the Cumberland field is 3000 tpd. The field is highly inclined, inclination varying from 12° to 32°. Overlying beds consist of shales and massive sandstone lenses of extreme toughness and are responsible for bumps when the stresses are relieved by extraction. At greatest depth these are among the deepest coal workings in the world. Depth of cover ranges from 2300 to 4000 ft. This prohibits room-and-pillar working and necessitates longwall operation. Working of contiguous seams concurrently to maintain output increases the already difficult conditions. The Sydney field, with a frontage of about 30 miles, is the most important of the Nova Scotia coal fields. With the exception of one small area it is now wholly submarine. Output is approximately 21,000 tpd. Seams are 21/2 to 8 ft thick, and cover in the areas varies from 600 to 2300 ft, with an average of 1500 ft from sea bottom to the top of the seam. The seams dip in a seaward direction, pitches ranging from 6" to 4.0". The shoreline is the last place of entry to the seams and distance from the bank to the working faces is generally over 3½ miles, in some cases as much as 6½ miles. Ventilation is a problem and requires the construction of large permanent airways. Getting and Loading Coal: In 1925, in view of heavy pressures exerted by thickness of cover overlying the seams, roadways and pillars of the room-and-pillar system being worked began to break up and coal was lost. It was decided that a change in the method of extraction in areas with heavy cover was necessary, and experiments were made with many short walls and longwalls varying between 90 and 250 ft. Trial and error proved that the best operating length was between 400 and 500 ft, delivering all coal to the dip, with roadways to the face following a level course. The change-over was gradual, and the technique of roof control developed with the system, so that falls to the face are now very infrequent. An advantage of longwall mining is that it yields 95 pct extraction, especially important in coal seams of high quality or in seams where faults or disturbances restrict the workable areas. This percentage of extraction is based on the fact that the longwall advancing system takes development faces where pitch permits instead of driving headings and leaving roadway pillars. This system yields high tonnage during development and limits loss in extraction to duff left during operations. Accompanying disadvantages, on the other hand, are the heavy construction cost of main roadways and the necessity of driving all new flank face roadways through the gob. If the main roadways are driven through the solid, and large enough pillars are left on each side for protection against flank face weights, then the width of solid coal is approximately 1700 ft. This represents 10.6 pct of the coal available, or 89.4 pct extraction of the whole. These pillars, however, are of such size that they provide a useful pillar drawing area as a final operation of the mine. It may be that although the seam cannot be generally mined by the room-and-pillar method, it can be adapted to

Jan 1, 1955

By M. P. Tixier, A. Poupon, M. E. Loy

Simple qualitative methods are explained for identifying those shaly sands in a well that are most likely to contain oil. A need for more precise measurement of the variables that enter shaly sand analysis is indicated. Field examples are given to illustrate the methods. A theoretical discussion of the quantitative interpretation of shaly sands is given as a basis for discussion and as a guide for the future. While not generally capable of practical application at the present time because of the lack of sufficient accuracy of the electric log data, these methods may become more feasible in the future as the result of the improved logging methods now being introduced. INTRODUCTION Experience has shown that for porous formations containing only a negligible amount of clayey material, reliable information on the fluid saturation and porosity of the reservoir rocks can usually be derived from the electrical logs. The interpretation is based on empirical formulae relating the true resistivity of a porous formation to its lithologic character, to the resistivity of the interstitial water, and to the proportions of water and hydrocarbons in the pores.' If the resistivity of the interstitial water is not known, its approximate value can be derived from the SP curve.2,3 The porosity can usually be determined to a good approximation from a MicroLog, or a MicroLaterolog.4,5 When the reservoir rocks contain an appreciable percentage of clayey material, an additional factor is introduced into the analysis. In a clean formation, the matrix is an electrical insulator, so that the ability of the formation to conduct current is due only to the conductivity of the electrolytes in the pores; in a shaly formation, the shale constitutes a part of the rock matrix able to conduct current, and influences the resistivity of the formation. On the other hand, the spontaneous emf's and the circulation of the SP currents are also responsive to the presence of the shaly material. All other factors being the same, the deflection of the SP curve opposite a shaly permeable formation is smaller than the deflection which would be observed if the formation were clean. The effects of the presence of shale on the resistivity and on the SP curves in a shaly formation are functions of the amount, of the specific resistivity, and of the geometrical distribution of the shaly material within the formation. The presence of shale therefore constitutes an important complication in electrical log interpretation. It is easy to realize, for example, that the presence of a large amount of clayey material in a sand may have such a predominant effect that the resistivity of the sand will not change very much, whether the pores are entirely filled with connate water, or contain a high saturation of hydrocarbons. Important research projects have been undertaken to investigate the influence of shales and to find a way to allow for this influence in the interpretation of a log. The problem of the SP in shaly sands was analyzed mathematically by H. G. Doll"' for the case of sands with shale laminations. Formulae expressing the resistivity of shaly formations were proposed by Patnode and Wyllie8 and de Witte.9 Other quite interesting theoretical and experimental results have been published by various authors.l0,ll,12,l3,14 The purpose of the present paper is to discuss an approach to electrical log interpretation in shaly sands. This approach is based partly on some of the theoretical and experimental data available in the literature and partly on observations made on actual field logs. Comparatively simple methods for qualitative and semi-quantitative analysis are described. These methods depend principally upon information provided by the electrical logs.

Jan 1, 1955

By John D. Wiebmer

Lignite Development in North Dakota is a "shotgun wedding" according to former state senator Robert L. Stroup-the unwilling groom (North Dakota) is being led to the altar by the nation's demand for increased coal production. The state's attorney general, Allen I. Olson, reacts to the situation this way: "We have no specific coal policy. We plan to keep the whole matter down to a dull roar and drag our feet until we see how we fit into the national energy picture." But despite a state administration labeled as "cautiously negative" toward resource development, North Dakota has experienced a twofold increase in annual lignite production within the last two years. And within the next four years, three more mine additions will double the state's output again.

Jan 8, 1977

By J. J. Scott

INTRODUCTION Rock mechanics theory needs to be applied to solve ground control problems, but in the final analysis the mining method and equipment which is used will be those which solve problems in a practical manner. Equipment and its inherent productivity and safety as supplied as "off the shelf" items by manufacturers often dictate the size of mine openings. Operators may tolerate falls of ground and associated hazards if productivity and safety problems are not too severe. In this whole process the costs of mining finally dictates just what system will be employed and what equipment will be used. True solutions to ground control problems sometimes require that the mine operator step back and totally rethink his problems. The true cost of accidents are often missed and hidden in overhead costs. Losses in production and accompanying profit are not fully considered, false economics, such as using inappropriate roof support devices, are undertaken at times when using the most expensive devices to totally eliminate accidents may be the best overall solution. Automation to remove the worker from the hazard area is often over- looked as the initial capitol expense to apply it may be high. Cooperation between mine engineers, mine operators, safety personnel, equipment suppliers, and government inspectors, both state and federal, is required to reach workable solutions that best meet the goals of all parties.

Jan 1, 1983

By Robert Bosworth

THE liability for damages to the surface caused by subsidence is an ever present threat in all underground mining. In ordinary lode mining, this threat rarely materializes into an action, due to the method of mining metaliferous ores, the dip of the veins, and the nature of the ground itself. In the mining of coal, however, this threat is very real. The coal frequently lies close to the surface, in beds underlying large areas of land the surface of which is suitable for agricultural or urban development. Furthermore, the coal is generally mined by the "room and pillar" method and when the pillars of coal are removed, the subjacent support is destroyed. This question of the duty of surface support has now become a very live one in Colorado. What was waste prairie land a few years ago, has become valuable agricultural land dotted here and there with growing towns. Frequent references. will be made here to Colorado conditions and adjudicated cases, but those conditions and cases and the conclusions to be drawn therefrom are equally applicable to all coal-mining states. Briefly stated, the question: is "Does the owner of the substrata or underlying mineral owe any duty of surface support to the owner of the superjacent strata or surface?" There is no absolute duty of support owed by the mineral owner and this assertion is made with full knowledge of numerous expressions to the contrary by courts of Pennsylvania, Illinois, Colorado and other coal-mining states. Whether or not any duty exists depends upon the circumstances surrounding each case, and since in the nature of things those circumstances are rarely identical, there can be no absolute rule.

Jan 1, 1927

By E. G. Leonardon, Irving B. Crosby

ElectRical prospecting by potential methods has been applied to mining problems for some years and determinations of the depth to bed rock have been made, but so far as is known it has not been used previously in the solution of civil engineering problems, such as investigating the foundations of dams. During the past winter and spring the senior author has been engaged as consulting geologist in studying the site of a large dam on the upper Connecticut River and has had the satisfaction of introducing electrical prospecting to foundation problems. The valley across which the dam is to be built is underlaid by a deeply buried preglacial gorge which is filled with glacial deposits. An extensive drilling campaign was in progress, but the work was unusually slow and expensive on account of the number of boulders in the overburden. The necessity of speeding up the work was pressing and it was decided that much of the needed information could be obtained by geophysical methods. After investigating the various methods and conferring with the company engineers and with the consulting engineer, Albert Crane, a contract was made with the Schlumberger Electrical Prospecting Methods, and an observer with apparatus was sent to the dam site the first of April, 1928. The problem was to determine the depth to rock at various points in order to prepare a contour map of the bed-rock surface and outline the buried gorge in the vicinity of the dam site. Over one hundred determinations of depth have been made and the bed-rock topography is now known in considerable detail. In addition, some information has been obtained about the nature of the overburden. The drilling has been continued; in fact, two additional drills have been put to work, and the electrical determinations have been checked at several points. At one of these the depth to rock given by the electrical work was 142 ft. and the drill found rock at 147 ft., an error of less than 4 per cent. The electrical work has also been checked by the geological study, and it has been found that when the electrical determinations are in agreement with the geological expectations they are probably reasonably accurate. When the electrical interpretation of the subsurface topography is incon-

Jan 1, 1929

By George W. Becker, Blair R. Benner, Ronald J. Roman

A computer model for the leaching of oxide copper heaps is developed and verified in laboratory tests. The model can be used to scale-up laboratory test data and to simulate the results of various operating philosophies on heap leaching. The model is applicable to leaching processes where the rate of leaching is controlled by a deficiency of reactant or diffusion of a reactant through the leached rock. Several examples are presented showing the sensitivity of the recovery vs. time curve to the variables in leaching.

Jan 1, 1975

By E. L. Brack

ASARCO's copper smelter at Hayden, Arizona has been upgraded to meet the increasing Environmental Protection Agency (EPA) and Occupational Safety and Health Administration (OSHA) environmental standards. Fortunately, means were found to accomplish these standards and provide for more efficient operations. With the adaptation of oxygen flash smelting and other related operations at Hayden, ASARCO has turned to the use of minicomputers and microprocessors for inter-plant communications and local area process control. The applications of distributive and computer controls to the plant operations at the Hayden smelter are discussed in this paper.

Jan 1, 1984

By P. H. Dudley

The present paper was suggested by the paper read by Mr. R. W. Hunt at the New York Meeting, in February last, on rail-sections (Trans., xvii., p. 778)) in the discussion of which I brought forward some of the facts and observations here stated. Professional duties having prevented me from writing out my part of that discussion in time for its publication in the volume containing Mr. Hunt's paper, I take the opportunity to present it in separate form and greater fulness. The striking features of the series of rail-sections submitted by Mr. Hunt are their shallow heads and heavy bases. In these respects they are quite similar to the section of the 80-1b. rail, which I designed in 1883 to meet the requirements of the traffic of the New York Central and Hudson River Railroad. Before the adoption of the 80-1b. rail, this road had modified its 65-1b. section from a shallow head with rounding sides and quite heavy base to one having a deep head and thin base. All the leading railroads had changed from the earlier sections of steel rails to those having deep heads and thin bases, conforming to the recommendations of the Committee on Rails of the American Society of Civil Engineers. It is but proper to say that the committee's investigations upon the wear of rails were made upon the rails manufactured under a practice quite different from that which obtained after the committee's recommendations were carried into effect. The results in actual service have been contrary to what was expected. Fig. 1 shows my 80-1b. section of 1883. Before designing this rail I had taken, at different times, a series of diagrams of the condition of tracks, with my car, on the New York Central and Hudson River Railroad, all of the New England railroads, and all but one of the trunk lines. My diagrams gave me an opportunity which could not otherwise have been obtained to see the wear of great numbers of sections of rails under widely varying conditions of traffic.

Jan 1, 1890

By Glenn C. Waterman

THE diamond drill is a very important tool in exploration and development testing and its use is increasing. In almost all cases results of diamond drilling are analyzed on the basis of grade and tons. A proper evaluation of core and sludge assays is important if drilling results are to be acceptable as a basis for geologic and engineering appraisal. The relatively wide variation in assay averages as calculated by various well-known combining methods indicates that the engineering choice of a method may affect the outcome of the drilling in terms of ore and waste. The problem of combining assay results from core and sludge samples has been discussed many times in conference and in the literature.'-' Most writers agree that the field of disagreement in methods is large and that the engineer on the job must consider features unique to his drilling, pick one of several combining methods, and depart from the rules when abnormal results come in. All the discussion to date can be summed up by the admission that as yet there is no fairly simple, generally acceptable combining method that is practicable over a wide range of drilling conditions, ground conditions, and ore occurrence. The combining problem is important in evaluating drilling results at Chuquicamata. Recently a reappraisal has been made of recovery variables and their effect on assays, with the result that a new combining method is offered which fits average drilling conditions and is mathematically reasonable. It is simple in application, fundamentally correct, and an improvement over most combining methods. At Chuquicamata diamond drillholes are used to outline the grade and position of blocks of normal and marginal grade oxide, mixed, and sulphide ore. Most holes penetrate all classes of material (and waste), and it is important for mining programs as well as ore reserves to know almost precisely the soluble and insoluble copper content of mineralized ground. At present three classes of ore are mined and treated differently. For an orderly sequence of mining operations which can provide regular daily tonnages of all three ore types and keep grade at certain levels with minimum variation, ore type and its grade must be predicted. Diamond drilling plus geologic mapping and bench sampling are tools for prediction. And drilling data are largely used to calculate grade of material more than a few meters away from bench faces. The orebody at Chuquicamata" is criss-crossed by millions of barren or mineralized weak to fairly strong slips and fault fissures. Mineralization is diverse and encompasses many quartz and oxide or sulphide-bearing copper veins, as well as seams and disseminated grains. Copper occurs in oxide or sulphide minerals or mixtures of these two mineral types. Rock conditions vary from intensely seri-citized (soft and porous) through clay-altered ground to almost fresh granodiorite. The result is an orebody which offers many obstacles to good and consistent core recovery in diamond drilling. Recovery varies considerably in the several alteration zones, the various types of oxide and sulphide ores, and the position and inclination of the drillhole within the complex fracture pattern. As core recovery drops sludge samples must be used with core samples to calculate grade. Many years of drilling at Chuquicamata indicate that in good grade oxide zones and in the sulphide areas core recovery is good and the ground uniformly mineralized. Moderate loss of core, therefore, does not markedly affect grade calculations based on core assays. An early core-sludge combining method used core assays at face value as indicating grade down to 50 pct core recovery, but below this recovery percentage sludge samples were used and weighted according to the standard Longyear chart. This method apparently did not introduce serious errors, but it abruptly used sludge assays with high weighting factors representing 100 pct return irrespective of actual percentage of sludge recovered. Recent drilling activities have been directed toward outlining the marginal ore areas. The non-uniform mineralization and generally poorer core recovery in such ground indicated that a more exact core-sludge combining method was required to equate wide differences between core and sludge assays and recovery. In fringe ore areas at Chuquicamata core recovery averages about 50 pct, from a minimum of 10 to 15 pct to a maximum of 100 pct. Sludge recovery is likewise variable and averages perhaps 80 pct even though holes are cemented as water return falls off. In a homogeneously mineralized area cut by many slips and faults, with hard and soft ribs, loss of core is loss of ground which has a grade similar to that of core recovered, and core assays approximate true grade. In this case sludge samples need not be used. However, it would be unusual to know beforehand that an area is uniformly mineralized, and in fact this condition is probably uncommon. Generally the distribution of valuable minerals in the ground does not exactly compare with their recoverability in core. Thus, in the usual case, loss of core decreases the

Jan 1, 1956

By Philip W. Henry

There is a difference of opinion among engineers on the subject of depreciation in general, and still more on its application to any given case The committee which was appointed by the American Society of Civil Engineers to make a report on Valuation of Public Utilities, states "There is no subject connected with valuation about which there are more diverse views than those relating to depreciation;" and in the discussion which took place upon the presentation of the preliminary report of this committee in January, 1914, that part of the report dealing with depreciation called forth more discussion than any other. From this discussion it was evident that the word "depreciation" was not always used in the same sense, so it is quite important that some definition be adopted. As this paper relates to the proper method of treating depreciation by corporations or individuals operating oil properties, both of which must make a return of annual net income to the government in connection with the income tax, it will be in order to adopt the definition of the Internal Revenue Bureau. Under date of Mar. 29, 1910, the Commissioner of Internal Revenue of the Treasury Department stated "Deduction on account of depreciation of property must be based on lifetime of property, its cost, value and use." Again, on Form 1035, "Return of Annual Net Income" for miscellaneous corporations (Section 2, Act of Congress approved Oct. 3, 1913), a deduction from gross income is allowed for "Total amount of depreciation for the year." This item is defined more particularly as follows: "The amount claimed under Item No. 5 (b) for depreciation should be such an amount as measures the loss which the corporation actually sustains during the year in the value of buildings, machinery, and such other property as is subject to depreciation on account of wear and tear, exhaustion, or obsolescence. The amount taken credit for on this account in order to be allowable should be so entered on the books as to constitute a liability against the assets of the corporation. The amount claimed under this item should not cover losses in the value of stocks and bonds. Decrease in the book value of securities owned, so far as such decrease represents a decline in the actual value of such securities, should be deducted under item 5 (a) of the return. "Where depreciation of physical property is made good by renewals, replacements,

Jan 1, 1915

By Robert H. Richards

While yet in correspondence with the heads of the several mining schools for the purpose of bringing the details of my Address into the best form for publication, a severe illness overtook me, and before I had recovered sufficiently to turn my attention to matters of this character the Address had been printed in its imperfect form for volume xv. of our Transactions, and the type had been distributed. The incorporation of the later corrections, in the form of a Supplement, became a necessity. To simplify the corrections as much as possible to the reader, every paragraph which has been revised in any way is here reprinted in full.

Jan 1, 1887

By R. B. Schluter, M. M. Fine

The Bureau of Mines' Twin Cities Metallurgy Research Center has of late concerned itself with reduction-induration of iron ore. One development provided a flowsheet for the agglomeration of green concentrate pellets in a reducing atmosphere created by a solid reductant external to the agglomerate. The product pellets were well metallized and physically competent. Maximum recommended operating temperature was 2100º F. In these studies, it was observed that unflred iron ore pellets indurated in a bed of anthracite or certain lignite chars frequently acquired exceptional crushing strength. A diagnosis of the phenomenon revealed that sulfur in the fuel was the probable cause. Accordingly, a research program was established to test this premise and to put it to practical use. Gaseous reductants were employed initially because of the ease and accuracy with which they could be metered.

Jan 10, 1969

By Clark A. Sumner, D. Arthur Burnham

A process for recovery of greater than 99% of the copper contained in acid leach solutions by sulfide precipitation using hydrogen sulfide as a hydrometallurgical reagent has been developed. The process incorporates the concept of hydrogen sulfide addition to recycled copper sulfide solids in a turbulent flow area where further precipitation or agglomeration occurs to increase the effective particle size of the copper sulfide precipitate. Because the recycled solids provide a larger number of particles upon which momentary excesses of hydrogen sulfide are absorbed, oversulfidization of the product is prevented. Copper sulfide precipitates produced in this manner in a laboratory test reactor showed greatly improved physical characteristics compared to copper sulfide precipitates obtained by conventional direct precipitation in which H2S is bubbled directly into acid leach liquor. Copper sulfide precipitates were evaluated and compared by use of filtration tests and settling rate studies. Electron microscopic examination of the solids confirmed the presence of larger particles in reactor-precipitated copper sulfide. Flotation was examined as an alternative to filtration for recovery of CuS precipitates and was found to achieve complete CuS removal without addition of collector or frothing reagents in a period of 2 min. In the same test reactor, a rapidly floatable CuS precipitate was also formed by mixing H2S with acid liquors used to leach crushed copper oxide-copper sulfide ore (LPF process). In the presence of gangue from the crushed ore, flotation of both natural and reactor-produced copper sulfides resulted in recovery of better than 90% of the copper contained in the crushed ore.

Jan 1, 1974

By Kendall Y. Keuhey

The Mercur Mine, developed and operated by Getty Mining Company, is designed to process 3,000 tons/day of gold ore from the historic Mercur Hill mining district. The mill process involves crushing, grinding, and cyanide carbon-in-leach for recovery of gold from ores ranging from 0.08 to 0.10 oz/ton. During the mining of normal grade ore, a substantial tonnage of material containing 0.02 to 0.05 oz/ton is mined and would be available for heap leaching. Also, Getty Mining is developing methods to test and prepare lower grade material for heap leaching by rejecting a coarse rock fraction from the grinding SAG mill circuit. This lower grade fraction is suitable for heap leaching, and its removal will upgrade the regular mill feed. Other test methods are directed toward leaching techniques on some 4.0 million tons of tailings that resulted from roasting and/or cyanide leaching Mercur ores in the early 1900's. The natural carbons associated with the Mercur deposits require special treatment to reduce their gold adsorption effects when leached with cyanide.

Jan 1, 1984

By F. W. Glaser

Metal-carbon-boron powder mixtures were hot pressed and the resulting specimens were studied by X-ray diffraction. It was found that regardless of the starting combination of the metal, carbon, or boron powders, a metal boride phase was always the major component in these samples. In the absence of carbon the boride phase formed on hot pressing depended only on the amount of boron present. Two new phases of the system Ti-B were found. They are Ti2B and Ti2B5. The existence of a controversial face-centered cubic phase of formula TiB was confirmed. Electrical resistivities were measured for various boride phases. It was found that the diborides are generally better conductors than the monoborides of the same metal. THE carbides and borides of the transition elements have very high melting points, in the range 2500° to 4000°C, and are therefore of interest as high temperature materials. The literature on the stability or chemical reactivity of these carbides and borides is very scarce. Various investigators'-" have demonstrated a relative instability of certain carbide phases in the presence of boron or boron-containing substances. In a recent publication, Glaserl demonstrated the stability of zirconium-boride (ZrB,) in the presence of carbon at temperatures in excess of approximately 2900°C, while during a preliminary investigation of boride phases, Steinitz' concluded that the diborides are stable in the presence of carbon while the monoborides of the fourth and fifth group are not, forming diborides plus carbides instead. Nelson, Willmore, and Womeldorph" have elaborated on the reaction B,C + 2TiC = 2TiB, + 3C, which was known to occur because of a relative instability of B,C and the great tendency towards TiB, formation at relatively low temperatures (approximately 1200°C). A similar study, involving as starting materials TiO, and B,C and resulting in TiB,, was recently described by Honak4, who observed the beginning of an exothermic reaction of a Ti0,-B,C powder mixture, which, when preheated in a hydrogen atmosphere to approximately 950°C, was carried to about 1600 °C by the heat of reaction. To shed more light on reactions of this type (Metal-C-B), the final product apparently always resulting in a boride phase at the expense of a carbide phase," a systematic investigation was started * Boride phases of various metals, as reported to date, are listed in Table I. and the following is an account of some of the results that were obtained. Materials, Preparation of Samples, Testing Methods The raw materials employed for this work consisted of various carbide, boride, and metal powders. as well as of boron and graphite powders. In cases where commercial grades of carbides were considered unsuitable because of low purity or excessive amounts of graphitic carbon, such carbide powders were prepared by this laboratory. The procedure for the preparation of carbide powders (zirconium carbide, titanium carbide, tantalum carbide, and niobium carbide) consisted of mixing graphite and the respective metal hydride powders in stoichio-metric proportions and subsequent heating of such mixtures in a hydrogen atmosphere in carbon crucibles. The heating was by high frequency to temperatures ranging between 1700" and 2100°C. The resulting carbide was then comminuted and screened to the desired particle size. ZrB, and TiB, powders were produced by the electrolysis of fused salt baths, according to the method described by Andriex.. The borides of niobium, vanadium, tantalum, molybdenum, chro-ium, and iron were obtained by mixing the respective metal and boron powders in the desired proportions. Such metal-boron mixtures were heated in a high frequency furnace to form boride powders. For each metal-carbon-boron group (Tables I1 through XI) a metal, its hydride, carbide or boride were mixed with carbon, boron or boron carbide powders. The additions of carbon, boron or boron carbide powders to any of these metals or metal compounds were calculated to satisfy a particular carbide or boride phase that according to the literature (Table I) had definitely been established by X-ray diffraction work. Samples of powder mixtures were hot pressed in graphite molds that were heated by direct conduction. The specimen dimensions were approximately 2.5X1X1 cm. Hot pressing temperatures were measured optically and maintained for approximately 30 sec under a constant pressure of about 1.3 ton per sq in. Wherever possible, an attempt to obtain maximum specimen density was made by temperature variation. Electrical resistivity testing was done by measuring potentiometrically the voltage drop over a length of 1.5 cm for a current of 10 amp, at room temperature. To obtain electrical resistivities for specific carbide or boride phases, values were plotted as a function of the respective sample densities

Jan 1, 1953

By Charles J. Haas

The resistance which rock bolts offer to shear movements along fracture planes and bedding planes near a mine opening has been suggested as one of the mechanisms by which bolts contribute to the support of weak rock in the roof or sidewall. That such displacements do occur is a fact fairly well known. Inspections of drill holes that have been drilled for some time often reveal that the holes have been offset at one or more points some time after drilling. Resin grouted bolts are often cited as being better able to resist shear movement in the roof because the annulus around the steel bar is filled with grout, whereas the sides of a conventional bolt will not contact the rock until a rather large shear displacement has occurred. When this project began, there were no experimental data to support or refute the above arguments. A laboratory shear test program was, therefore, devised to evaluate the effectiveness of various types of bolts in resisting shear under various conditions which exist in situ. Research reported earlier (Haas, 1976) included a theoretical analysis of a conventional bolt intersecting a shear surface, and experimental data on the shear resistance of smooth (lapped) shear surfaces, intersected at various angles by conventional and resin grouted bolts. The present paper extends the previous analysis to consider the effect of the initial roughness of the shear surface. Joint roughness is a particularly significant factor since undulations of the shear surface cause dilatation of the joint, which in turn, causes high tensile loading of a bolt intersecting the joint.

Jan 1, 1982

By F. L. Fayers, J. W. Sheldon

It has been a matter of concern to the petroleum industry to determine what effect the capillary term has on saturation profiles, since these profiles determine ultimate economic oil recovery. In their original solution of the non-capillary two-phase flow problem, Buckley and Leverett&apos; showed that solutions of the two-phase flow equations became multiple-valued in saturation. Since it is physically unrealistic for saturation to have more than one value at a given position, Buckley and Leverett resolved this difficulty by introducing a saturation discontinuity or "shock". They evaluated the strength and position of the shock from material balance considerations. Recently Sheldon, Zondek and Cardwell&apos; have demonstrated that the method of characteristics can be used to solve this and related problems. Owing to capillarity it is physically unlikely that saturation shocks occur in an oil reservoir. Buckley and Leverett suggested that the effects of interfacial tensions should be included in the fluid flow equations, if the solutions are to give continuous single-valued saturation distributions. An analogy to this problem is encountered in the theory of wpersonic compressible flow where the introduction of the viscous term eliminates shock behavior. However, the mathematical form of the capillary term in the porous medium fluid flow equations differs from the viscosity term in the Navier-Stokes equation. This paper demons~rates that capillarity does eliminate the triple-valued behavior and indicates the magnitude of capillary and gravity effects on the saturation profile for a variety of systems. DERIVATION OF LAGRANGIAN AND EULERIAN FORMS OF THE DISPLACEMENT EQUATION The equations representing capillary, two-phase flow are given by applying Darcy&apos;s law and a material balance condition for each phase. The familiar Darcy and material balance equations are as follows. ^U*^0......<3, ^^r=°......c« S is the saturation of the water phase, and B the inclination of the porous medium. Pressures in the oil and water phases differ by the capillary pressure associated with the interfacial tensions, P.-P.=pAS)..........(5) where p, is determined experimentally. The porous medium is said to be oil-wet if P, > Po and water-wet if P, < Po. If the fluids are incompressible and the total volumetric flow rate, q, is constant, then v, f v,, = q...........(6) By eliminating P.. /»,, and v, from Eqs. 1, 2, 5 and 6 it follows that It is convenient to normalize the variables in the problem after the manner of Handy and Hadley% so that Eqs. 3 and 7 become where j,, = —, the fractional flow of water; (F, is a capillary pressure normalizing constant.) Substitution of Eq. 9 into Eq. 8 gives In this partial differential equation for an initial value