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By C. H. Bushell, M. Malnarich

REAGENT control in flotation is more an art than a science. Operators vary the amount of reagents used according to the metallurgy obtained. The amount of collector may be increased, for example, if tailings losses are high and decreased if concentrate grades are low. Often control decisions are based on only one assay, such as the tailings analysis by microscopic or quick assaying techniques. This method of control has given reasonably satisfactory results but is clearly far from perfect. No adjustment in reagent addition can be made until information is received that metallurgy has deteriorated, that is, until the optimum time for adjustment has already passed. The first adjustment probably will be an overcorrection or an undercor-rection, and time will be required to establish the best new addition rates. The seriousness of the problem depends on variability of the ore. In ore of continually shifting composition and complexity, mill results will shift back and forth from excessively high concentrate grades and high tailings losses to excessively low concentrate grades and low tailings losses. Economic incentive for improved control also increases with increasing size of mill. Even a small improvement in percentage recovery in a large mill pays for expensive control equipment. The Sullivan concentrator of The Consolidated Mining & Smelting Co. of Canada Ltd. at Kimberley, B. C., treats 11,000 tpd of complex lead-zinc ore, a scale of operations which justifies an extensive research program on automatic control. Work has followed two separate lines: 1) development of faster analytical methods for pulps, and 2) study of metallurgical control by measurement and control of reagent concentrations in flotation solutions. Although work is incomplete in both phases, a progress report can be given on the development of reagent control. The primary requirement in flotation is selective attachment of particles of one mineral to air bubbles while the other minerals remain detached. Hydrophobicity of mineral surfaces is established by the collector, which coats the minerals by adsorption. The requirement of selective affinity for air bubbles is therefore equivalent to requiring selective adsorption of a collector. Adsorption theory, supported by abundant experimental evidence, shows that the amount of a reagent adsorbed at a surface must be a function of its residual concentration in solution. As the concentration in solution rises, the amount adsorbed increases, usually following some type of logarithmic curve. Selective adsorption must depend on a natural difference in the minerals, but ordinarily the differences in affinity are not great enough to obtain complete coverage of one mineral while the others are completely uncoated. An optimum concentration will exist, below which coverage will not be great enough to float enough of the valuable mineral, and above which too much unwanted mineral will float. All other reagents, except some frothers, affect the amount of collector adsorbing at constant residual collector concentration in solution, either by adsorbing competitively with the collector, or by a

Jan 1, 1957

"The richest hill on earth" had its start as a placer gold camp in 1864. The gold was quickly worked out but then silver was discovered in the copper ores in Butte Hill. After ten years as a silver camp, rich copper ore was found in 1881 and the future of Butte was assured. In the early 1960's, the Anaconda Company began the shift from underground mining at Butte to open pit mining of the enormous, low-grade copper deposit surrounding the old high-grade veins which had been worked for over eighty years. Their first primary crusher was a 66 in. by 84 in. (168 m by 2. 13 m) jaw crusher. The pit was expanded at such a rate, however, that its location became endangered. Additionally, the new Clyde Weed concentrator had been built at Butte; and in 1969 construction of a new gyratory installation was started. The primary crusher at the Clyde Weed concentrator is one of the most unusual primary crushers ever put into operation. At times, the ore from the Berkeley pit contains large amounts of timber, pipe, rails, and so forth left in the ore body from the old underground mining operations; and the crusher installation had to be designed so this tramp material could be readily removed ahead of the crusher. Ore is delivered by 100- and 150-ton trucks to two side-by-side dump pockets, each with a capacity of 400 tons. Originally, a hydraulic grapple was installed on the dividing wall between the dump pockets; but it was found that oversize boulders large enough to block the discharge opening at the bottom of each hopper were rapidly covered with other material, and the grapple was unable to reach and remove the large boulders causing the blockage. In addition, the grapple was located in a quite vulnerable position; and it sustained damage from being struck by rock dumped from the trucks. Oversize boulders are now blasted. Underneath each pocket is a vibrating grizzly feeder with the grizzly bars set at 4 in. (102 mm). These feeder-grizzlies are 84 in. (2. 13 m) wide and 32 .ft (9. 75 m) long, with grizzly sections 16 ft (4. 88 m) long. Approximately 75% of the ore passes through the grizzlies and grizzly oversize then drops onto two parallel 6 ft (1. 83 m) wide picking belts. The belts are 3 in. (76 mm) thick and rest on idlers supported by air- filled rubber shock absorbers. At the side of each picking belt is a hydraulic grapple which is used to remove timbers too large to be removed by hand. At times, four men are used to remove timber and tramp iron. Timbers are dropped onto a transverse belt conveyor which discharges into a truck stationed outside the primary crusher building,

Jan 1, 1978

By Jon J. Connor, Richard J. Ebens

The trace element geochemistry of the cherty clay-rich carbonate residuum cover in much of southern Missouri and adjacent parts of Arkansas was investigated during the earl 1970's as part of a statewide geochemical survey of Missouri.1-8 The study was environmentally oriented and focused on the character of the "natural" geochemical environment of the state. A great deal of effort was made to collect samples that would represent the geochemical variability of the near-surface environment in an unbiased way. In general, this was accomplished by use of randomization procedures in both collection and analysis. Most of the sampling sites were selected randomly to prevent personal bias from entering into sample selection. In the laboratory, samples were analyzed in a randomized sequence to circumvent the effects of any systematic bias that may have entered into the laboratory procedures. Several residuum samples contained 1000 parts per million (ppm) or more of lead, zinc, and barium. All these samples were collected in Washington County in or near the barite district. The frequency distributions shown in [Fig. 1] indicate just how unusual the residuum in Washington County is with regard to these three metals and copper. The shaded parts of the histograms represent the samples from Washington County. The histograms are based on analyses of 141 samples from 38 localities scattered over the subcrop of the Cambrian and Ordovician formations of southern Missouri and northern Arkansas. Eight localities were selected randomly from over each of four bedrock units-Bonneterre dolomite; Potosi and Erninence dolomites; Roubidoux formation; and the Jefferson City, Powell, and Cotter dolomites. Six localities were selected randomly from over the Gasconade dolomite. The geochemical "background" for these elements in residuum is defined by samples from outside of Washington County (unshaded part of the histogram); lognormal distribution curves have been fitted to the corresponding parts of each histogram. Clearly, these lognormal models (curves) describe well the background distributions of zinc, barium, and copper. The background histogram of lead concentrations is notably skewed to the right compared to the model. Lognormal distribution curves were also fitted to the data on lead, copper, and zinc in the residuum of Washington County. No model was fitted for barium in Washington County because of the distinct bimodality in the data. All samples containing more than 1000 ppm were collected in the barite mining district and obviously reflect the elevated levels of barium in the district. The frequency distributions for Washington County are clearly anomalous with respect to the background data, those for lead, zinc, and barium being exceptionally so. The models for anomalous lead, zinc, and copper seem to fit reasonably well, although the scatters in the lead and zinc anomalies are considerably larger than the scatters in the corresponding background data. The copper anomaly, although probably real, is greatly subdued compared to that of lead and zinc. Also, it has a scatter similar to that of the copper background, rather than being enhanced like the scatters of the lead and zinc anomalies. On the basis of techniques described by Miesch,4 the chemical abundance of copper, lead, and zinc in residuum in

Jan 1, 1980

By C. S. Matthews, H. C. Lefkovits

The purposes of this study are (a) to determine the accuracy of a previously proposed method for calclrlating average reservoir pressure and (b) to find a method for estimating the shape of the drainage area of each well in a bounded reservoir. Both numerical and experimental methods are used to determine the pressure distribution and the location of drainage bounrlaries in idealized reservoirs at steady state. A number of photographs are included which show flowlines and drainage areas in reservoir models of various shapes. Results from these studies are then compared with those from the previously proposed method for calculating average reservoir pressure.' It is concluded that the previously proposed method is sufficiently accurate for most reservoir engineering purposes. In addition, some simple methods are given for estimation of .shapes of drainage boundaries. INTRODUCTION A method for the determination of the average Pressure in a bounded reservoir was previously developed by Matthews, Brons, and Hazebroek,1 and presented at the Petroleum Branch Fall Meeting in 1953.† In aP- plying this method, the reservoir was first divided into drainage volumes; second, the average pressure in each drainage volume was calculated; and third, these drainage volume pressures were averaged volumetrically to give the average pressure in the reservoir. It was shown that each well's individual drainage volume, at steady state, is proportional to the well's production rate; however. the question of how to estimate accurately the shape of these drainage volumes could not be answered. Furthermore, in calculating the average pressure for each drainage area, the actual drainage areas, which are in general quite asymmetric in shape, are replaced with one of the symmetric prototypes given in T.P. 3876. A question may then be raised as to how accurately the symmetrical figures depict the pressure behavior of the actual unsymmetrical drainage areas. The study reported on in the present paper was made in an attempt to shed light on the above questions. METHODS It was shown in T.P. 3876 that after a period of steady rate of withdrawal the rate of pressure decline becomes constant at every point in a bounded reservoir containing a single-phase fluid whose compressibility is small and constant and whose viscosity is con-

Jan 1, 1956

By Richard Moldenke

Note.—At the request of the American Society for Testing Materials, this paper wau presented at the New York Meeting, February, 1923, in connection with the "A. S. T. M. Tentative Specifications for Foundry Pig Iron" and The Physics of Cast Iron and Its Bearing on All Cast-iron Products and Specifications for Cast Iron," by William R. Webster. The discussion of this paper was not published by the Institute, having been referred to the American Society for Testing Materials, as these papers were again presented for discussion at the meeting of that Society at Atlantic Cit;, June, 1923. The result of these discussions is shown in the following abstract from the Committee A3 report on cast-iron specifications: During the past year, the adaptability of the tentative specifications for pig iron for the steel and the malleable foundry, in addition to the iron foundry proper, have been studied and suitable changes and rearrangements are being made which appear to cover the situation properly. In spite of the experience of foundrymen to the contrary, there is a persistent call from consumers of castinga for defixute compositions for given lines of castings. To respond to this call for information desired by the consumer—information every modern foundryman should have and guide his procedure with—the committee has requested the Sub-committee on General Cast!rigs to take up this matter with a view of constructing tables of analyses rang? for Important lines of foundry product. These analyses ranges however, are to be given as "recommendgd practice" and not intended for actual spkcification. In this way, the consumer of the foundry product will get the information he wants; the furnace-man will be gulded in the production of his pig irons to prepare for analyses ranges which are more closely standardized and have less of "freak" requirements made by Inexperienced metallurgists; and the foundryman will not be bound by chemical requirements he knows will not work out well with his particular reglonal supphes of melting stock. Doctor Moldenke's paper is published in full as it shows clearly the good results of the Institute's work in 1895 and a comparison of the same with the present knowledge of the metallurgy of cast iron—Bradley Stoughton, Chairman, Iron and Steel Committee. In MaRch, 1895, William R. Webster contributed a carefully prepared scheme for the study of the physics of cast iron to the American Institute of Mining Engineers at its Florida meeting. This summary of the interrelation of the chemical and physical characteristics of cast iron with the operations of daily foundry practice caused a wide discussion of the then little understood though universally used material. Mr. Webster has requested the author to bring the subject up to date and, while glad to comply, he must pay that with all the knowledge that has been gained in the quarter century that has passed since the subject was brought to the attention of the Institute—and we now are pretty familiar with cast iron in all its aspects—there is practically noth-

Jan 1, 1923

By John T. Galey

An intensified search for gas, because of the shortage in the area, led to the completion of 482 wells in southwestern Pennsylvania during 1944. This number was 27 more than the number in 1943, but 118 of the wells were nonproductive. Shallow Development Oil The initial production of the 25 producers completed was slightly less than was obtained from four wells above last yearJs total. The largest well, 20 bbl., was found in the Big Injun (Miss.) sand in Morris township, Washington County, at a depth of 2085 ft. It is currently producing about 3 bbl. per day. Outside the main producing area, a well drilled in Nicholson township, Fayette County, obtained 15 bbl. per day from 1864 to 1866 ft. in the Gantz (U. Dev.) sand. The Federal subsidy of 75¢ per barrel, effective , 1, 1944, which increased to $3.40 the current pipe-line price in this area, seems to have been insufficient stimulus to materially increase drilling for oil, although more repressuring operations are contemplated. Gas The volume of initial open flow of gas developed is off slightly over 10 million cu. ft. from 36 more wells than were completed a year ago, and the average rock pressure of this new gas is more than 100 lb. less than that obtained a year ago. Apparently the best find of this year was reported in Patton township, Allegheny County, where an initial volume of over 4 million cu. ft, was encountered in the Speechley (U. Dev.) sand at a depth of from 3293 to 3303 ft. The rock pressure there, 1190 lb. per sq. in., is nearly virgin. However, this area is fairly well delimited by dry holes and docs not seem nearly so promising from the standpoint of reserves as the Armbrust area, which was discovered several Years ago. In the patton township locality another fair well with 1303 M cu. ft. open flow and 400 lb. Per 59. in. rock pressure completed in the Hundred-Foot (U. Dev.) sand at a a depth of 2125 feet. Armstrong County, as usual, leads in number of producing wells completed with lowest percentage of failures but ranks only fou th in volume of production developed. Only two wells here had more than one million cubic feet capacity, but the rock pessure was under 200 Ib. Per sq. inch. In Fayette County, where the largest portion of new gas was found, activity appears to be shifting eastward. In Luzerne township, four producers having a total combined initial open flow in excess of 10 million cu, ft, were found in the Big Injun (Miss.) sand. The best well was found also in the same area in the Speechley sand (U. Dev.) at a depth of 3460 ft. This was a 2500-M CU. ft. well in Georges township. Several wells of more than a million cubic feet capacity were found in Jackson, Washington, Wayne and Whitley townships, Greene County. None of these, however, was in a new pool.

Jan 1, 1945

By Donald C. Branford

The place of geophysics in the curriculum of a college or an engineering school has been much discussed. There is uncertainty as to whether the graduate may be called a "geological geophysicist" or a "geophysical geologist." Some have used the term "geophysical engineer." The question of the specialized undergraduate curriculum versus a general undergraduate training in the parent sciences with the addition of cultural electives has been and is still being argued. The literature during the past Io years shows that there are two rather definite schools of thought on this subject. The prospective employer, engaged in active geophysical exploration, visualizes the graduate in terms of the routine work he will do in the field. The school administrator, grounded in many years of tradition in the cultural type of education, visualizes the graduate in terms of his position in a political and cultural society as well as in his work as a professional geophysicist. Both are right and both wrong. A third school of thought, that of the student himself, has been almost totally ignored. It would seem best, for the purpose of this paper, to treat these various points and allied material separately, and then to integrate the conclusions as a set of conditions that should be fulfilled as far as possible in the ideal curriculum. At the end the relationship of our findings to the curriculum that has been tentatively adopted at the University of Pittsburgh will be considered. Divisions in Definition or Field of Geophysics Geophysics is that field of science in which geological and earth problems are solved with the use of physical and chemical techniques. Altogether too much attention has been paid to the artificial separation in this science of the commercial applications from the so-called "pure" or academic geophysics. The only real difference in the two types of work is one of scale of operations. In the commercial field the scale is large in terms of fmancial backing, immediate economic importance, and speed of operations. In the academic field the scale is large in terms of breadth of interests, in eventual economic importance, and length of time allowable for theoretical investigations. In exploration work the scale is small in terms of dimensions, since the methods are applicable only to the outer 20,000 ft. or SO of the earth's crust. In academic work, the student is not pressed for results of immediate economic importance, and therefore can afford to spend more time on problems that seem unimportant at the time. In the sense of background knowledge of the parent sciences necessary to the geophysicist, no essential difference can be found. Physics, chemistry, geology, and mathematics together, and in their proper proportionate amounts, form the matrix from which geophysical ideas and methods may spring.

Jan 1, 1946

By R. J. P. Lyon

It is usually assumed1,2 that ore deposition is relatively slow, taking place over tens of thousands of years. Yet many syntheses and phase changes can be completed in the laboratory in a matter of hours, under essentially the same pressure and temperature conditions to be expected in nature. Much has been written about geothermometry, and Ingerson's recent summary2 an hardly be surpassed. Rather than review the whole field, this article will deal specifically with diffusion rates in solid solutions* of the sulfide minerals. In the study and application of temperature-indicating phenomena to problems of ore genesis, the time factor has often been neglected. In the particular case of sulfide systems, the speed of unmixing and diffusion of the constituent phases in the solid solutions has seldom been considered when laboratory data are applied to the genesis of natural sulfide ores. Chalcopyrite and Bornite Intergrowths: Chal-copyrite and bornite are copper iron sulfides often found in association. In many ores the individual grains appear to have mutual boundaries, but bladed intergrowths are quite common. Early workers in the field, " deduced that while these patterns could have been formed by replacement, they could be experimentally reproduced from the unmixing of solid solutions of the two previously homogenized at 475°C. Much of the ore from Roan Antelope mine in the Rhodesian Copper Belt exhibits these bornite-chalcopyrite intergrowths, from which it has been deduced that the ores could have been subjected to a minimum temperature of 475°C. In several recent interpretations,' this temperature rise has been considered to result from a regional metamorphism of an original syngenetic copper-bearing sediment. The same temperature rise to 475 °C was suggested from similar intergrowths in the sulfide- • The general subject of solid solution in structurally similar minerels is treated in many texts'. :' ' and will not be further elucidated. bearing ore from the Happy Jack uranium mine8 on tile Colorado Plateau, although this temperature had a different origin. In 1931 SchwartzV etermined that 475°C represented the unmixing temperature of his bornite-chalcopyrite solid solutions but, more important, that the characteristic intergrowths of bornite and chalcopyrite could be preserved only if the solid solution were chilled quickly to room temperature. This significant corollary has been almost completely neglected by both Davis and Miller.7,8 By suitably decreasing the cooling rate, Schwartz found that the original pattern of the intergrowth was reformed (Fig. 1). The cooling rate was critical. In cooling from 475°C to room temperature in

Jan 1, 1960

By E. O. Bennett

When oil is takcn from a subsurface structure it is generally accompanied by gas. The gas thus produced represents the 1ighter hydrocarbons present in the original petroleurn accurnulation, which arc, insoluble at the reduced pressures created around a well when it is producing. In some cases the gas accompanying the oil may come from a gas-bearing stratum not in immediate contact with the oil. The amount of gas from the formation accompanying each barrel of oil is called the natural gas factor. Energy Causing Well to Flow The force causing the flow of fluid into a well when it is drilled into an oil-bearing structure is the pressure differential between the well bore and the formation. A pressure differential between the formation and the well may be set up in three different ways, as follows: (1) pressure due to gas in solution or in contact with the oil, (2) hydrostatic pressure due to water in contact with the oil (it may be either bottom or edge water); (3) pressure difference due to forcc of gravity. The greatest source of energy for producing oil from most oil fields, and the one over which the greatest control can be exerted, is the poten-tial energy in the gas under pressure in the formation. In any pool there is but a definite amount of gas available. Part of this gas may be in solution and a part in contact with oil. The amount of gas in solution in each bard of oil depends on the pressure and temperature in the formation and the characteristics of the oil. Whether the gas is in solution or in contact with the oil it is necessary to use the energy it, contains as efficiently as possible in order to obtain the greatest ultimate recovery from the pool. Bush1 states: Gas being the most efficient aid in moving the oil, it is important to prevent more of it from escaping than is necessary to remove the maximum amount of oil from the sands. The conservation of this energy is far more important to the producer of oil than his returns from sales of gas to distributing companies. The flowing life of the well is the most profitable period, and keeping the gas-oil ratio low, or keeping the

Jan 1, 1928

By Edward C. J. Urban

THE nature of certain pernicious dusts commonly encountered in the removal of iron ore from the underground mines of the Lake Superior district is recognized, and appropriate measures for the protection of individuals subjected to these dusts have been applied in most mines for more than to years. Undoubtedly the principal concern has been demanded by dusts that contain as a constituent silicon dioxide in its free or uncombined state, the mineral we all know as quartz or chert. Investigators1 agree that among the factors that determine injurious exposure are: the composition of the dust, the concentration suspended at the breathing level and the duration of the exposure. We have on hand data that describe the experience of several underground iron-ore mines in the Upper Peninsula of Michigan and Wisconsin in regard to these factors, which in a large measure also reflect the general experience of the other mines of the area. We believe that a discussion of these data will indicate the effectiveness of the dust-control methods that have been considered good practice and that have been employed over a long period of time. A silicosis-prevention program was ini¬tiated by the iron mines of the Upper Peninsula in 1934 and records of mine dustiness are available from that date to the present time. For dust conditions previous to 1934 the results of observations conducted in other industries can be utilized together with local information pertaining to the development and initial application of rock-drilling equipment. DEVELOPMENT OF ROCK DRILLS The earliest percussion rock drill that later had any industrial use was patented in this country in 1851.2 This machine was later modified and improved in many respects but its piston-type design remained standard for drilling equipment in most mines for the next 63 years.3 The piston drill was operated either by air or by steam. It employed a solid drill steel, fastened directly to the reciprocating piston. The drilling was done dry except in holes directed at an angle below the horizontal. For these water was manually injected into the hole and the pumping action of the rod expelled the sludge. The dry cuttings from drillings inclined above the horizontal fell from the hole by their own weight. A development of major importance from the standpoint of dust generation and control occurred with the invention by J. G. Leyner of a hammer-type rock drill that embodied the essential feature of a drill steel with a hollow core. His patent was secured in 1897. The machine was a radical departure in drilling method. The steel was held loosely in a chuck while

Jan 1, 1947

By C. F. Seaman, R. G. Clay

The mines of The Tennessee Copper Co. are in the Ducktown Basin, in southeastern Tennessee. The ore is a heavy sulphide consisting principally of chalcopyrite, pyrite and pyrrhotite and in places running as high as 95 per cent sulphur and iron. Operating mines are the Burra Burra and Eureka. Development ' of the Boyd ore body was started in 1940. Late in 1939 the decision was made to change from fuse and cap to all electric blasting and at the same time electric cap lamps were installed to replace carbide. These changes were made to improve the safety and working conditions in the mines. Both the Bureau of Mines and explosives manufacturers recommended electric blasting as a safety measure. Until that rime, electric blasting had been used only for shaft sinking and in places where it was absolutely essential for safety. Very few men had had experience in this type of blasting, as the places where it was required were those that also required direct supervision. It was necessary for the mining staff to work out methods, procedure, and equipment to be used. In this they were ably assisted by representatives of the explosives manufacturers. Many mistakes were made and it is the purpose of this paper to describe the original plan and installation, show the major difficulties encountered. the remedies applied, and present the practices now being followed. Comparison of costs of the two methods will also be discussed. About six months were required to make the change and eliminate fuse and cap blasting entirely. The first electric blasting stations were installed in November 1939 and the last fuse was taken underground in May 1940. A very unusual condition forced the use of fuse for a short time in one heading after May, but that will be discussed later. Original Equipment and Installation A power line is generally recommended as a source of current where permanent blasting stations are used. The plan worked out for electric blasting was to tap 220-volt alternating-current lines when they were available and use 250-volt direct current from the trolley wire at all other places. A blasting machine would be available at each shaft for places where there were no regular blasting stations and could be used if a regular station should fail. The equipment decided on was No. 16 duplex reinforced lamp cord for lines to within about 50 ft. of the working place and No. 20 lead wire from that point to the shots. An open double-pole, double-throw switch with slate base was put in the circuit, so connected that the down position would close the line to the shots 2nd the up position would connect it to the Iine from the power contactor. This switch was located in the circuit as close to the blast as was safe (should the shots explode when it was placed in the shooting

Jan 1, 1943

By W. A. Boyer

If the safety factor of a mine hoisting rope is checked for the lowest depth, is the rope then safe for all levels? The answer here is no. A new set of volues is proposed. HOISTS for metal mines are seldom designed for one particular depth. They are intended for an ultimate load and depth with a given speed but are first used to hoist from a shallower depth, gradually working to the ultimate as development progresses. Sometimes the mine is not worked on the upper levels because of the kind and grade of ore, the mining of which depends on market conditions. In this case the hoist may be used in the same working shift to lift ore, for example, from the 1200 and the 4400-ft level. This means that the safety factor of the rope should be considered not only when it hoists from the greatest depth, but also when it hoists from the upper levels. According to present safety factor standards, a hoisting rope can have ample safety margin when hoisting from the lower or lowest level and still be shy of sufficient safety margin when hoisting from the upper levels. This will be shown by curves presented in this paper. The safety factor of a hoisting rope as defined by the U. S. Bureau of Mines' is the ratio of the static load to the tensile strength of the rope. It is calculated by dividing the breaking strength of the rope* In Fig. 1 curves A and B are plotted against the vertical suspended distance, which is the abscissa, and the factor of safety, the ordinates. Curve C is a smooth curve drawn through the midpoints of the steps of curve A and would more closely represent the minimum factor of safety for new rope for any particular depth. Curve D is a smooth curve drawn through the midpoints of the steps of Curve B and would more closely represent the minimum factor of safety when ropes should be discarded for any particular depth. If a new rope were applied with a minimum factor of safety, then the vertical distance or ordinate under curve D for any particular depth, divided by the vertical distance or ordinate under curve C, would represent the minimum percent of remaining area intact that could be allowed before the rope would be discarded. Also, that section of the ordinate between curve C and curve D is proportional to the section of the area of the new rope that could be allowed for wear and broken wires. Considering that point at 2500 ft on the chart, the ratio do to co would be the minimum percent of rope area remaining at the point of discard. The ratio of that portion of the ordinate cd to co would represent maximum percent of area of a new rope available for wear and broken wires.

Jan 1, 1955

By A. D. Dorenfeld

Factorial design is a mathematical method of drawing valid conclusions from a series of tests made in a predetermined pattern. It is applied to flotation ore testing using, in this case, five variables, each varied twice, requiring a total of 32 tests. Representative conclusions are checked by repeat testing. THE factorial design method of testing is particu-J- larly applicable to mill-scale plant tests because 1—the conclusions take into consideration all variabilities of the plant, 2—there are no subjective guesses as to the conclusions, 3—the number of tests necessary to check large numbers of variables are at a minimum, and 4—interactions between variables can be discerned. It is equally applicable to laboratory testing because once the variables to be investigated have been decided upon, as well as the number of levels for each variable, a well-trained assistant can run the tests, since the test work follows a predetermined mathematical pattern, leaving the conclusions to be found by the person in charge. The conclusion takes into consideration all errors of the tests. To illustrate the use of factorial design, a sample of copper porphyry ore was tested, the copper content of the batches not being uniform. Five variables were investigated, each at two levels, requiring only 32 tests. After suitable arithmetical analysis, conclusions for copper recovery and concentrate grade are arrived at. Four cases of these conclusions are checked by repeat testing. Testing Methods Reproducibility of flotation tests varies widely, depending on the experimenter, the process, the equipment, and probably a host of unknown and perhaps interacting variables. If the reproducibility of tests is good, let us say to 0.2 pct recovery of a certain metal, and the minimum difference between tests upon the variation of the interested variables is, say 1 pct, then simple averaging and inspection is ample in interpreting results. However, often such a happy situation is not the case. This is particularly true in mill-scale operations where the ores are variable, and a host of minor operating fluctuations common to most milling, and familiar to all mill operators, is the normal state of affairs. Yet, mill scale tests are run, the results averaged, and conclusions drawn therefrom. The trustworthiness of such results should be measured by comparison to the experimental error—the measure of the degree of reproducibility of the results. This comparison is a test of significance. Thus, if daily results of recoveries average 80.0 pct, and a spread of 78.0 pct to 82.0 pct, is recorded, then, clearly, on changing mill operations, an 81 pct recovery cannot be construed as significant by using averages, since it is well within the previous test limits. By using rigorous statistical tests of significance, conclusions arrived at by inspection or simple averaging may be shown to be based on inadequate evidence or even wrong. In any event, a plant super-

Jan 1, 1952

By Charles C. Rodd

PRIOR to 1938, proration procedure in Kansas required the physical testing of wells in order to set up a basis for allocating production. Subsequently the use of liquid-level data and bottom-hole pressure data was authorized on certain types of wells as a basis for calculating well capacities, thereby permitting wells to be tested at low rates and with small volumes of production. This paper gives some attention to the equipment used for obtaining liquid-level data but in general it is devoted to the theoretical and physical factors involved in well testing by drawdown methods. These tests are divided into two classifications: (1) liquid-level data used directly to calculate capacities, and (2) liquid-level data and liquid-level measuring equipment used in connection with other data to determine bottom-hole pressure. Equations and calculations for both types of tests are shown. Well capacities calculated from fill-up data have not been used generally but offer a simple means of testing some wells. Theoretical factors and calculations for such tests are shown. IN the state of Kansas, proration laws required that the "ability of a well to produce" be given consideration in state allocation orders and for many years this was the only factor used in prorating the allowed oil. The "ability of a well to produce" was the quantity of oil that could be obtained from a well in a prescribed period, usually 24 hr. Early in the history of State-regulated proration, Kansas operators realized that, unless some restriction were placed upon equipment, the competition between producers to obtain large well "potentials" might result in a race to install the largest pumping equipment. Consequently, restrictions were placed upon the size of tubing and length of pump stroke that could be used in pumping wells during potential tests. Later, restrictions were placed also upon the size of tubing and chokes that could be used in flowing walls. Despite the restriction set-up, better pumping equipment was designed and installed, pumps were improved, and various devices were invoked to obtain high rates of production for short periods of time. By 1937 pumping cycles of forty to forty-five 54-in. strokes per minute were not uncommon, and numerous wells had produced at the rate of 3500 bbl. per day for short periods. Such production was far beyond any rate that could be long sustained. Equipment necessary to take "potentials" was several times larger than that needed for normal producing purposes, and repairs and replacements required by the high pumping speeds were excessive. Also, many operators were beginning to suspect that the high rates of production maintained during "potential" tests were conducive to early water encroachment. By 1938 large-scale physical testing of wells in Kansas ended, the use of the bottom-hole pressure gauge had become common, engineers and operators were becoming familiar with the term "productivity index," and bottom-hole pressure was being used as a factor in allocation in many areas. There were many drawbacks, however, to large-scale use of

Jan 1, 1942

By Anal J. Mehta, Theodore D. Tiemann

Growing demand for steel and increasing dependence upon foreign ore makes the need for developing the technology for treatment of low-grade iron ores greater than ever before• Representative of these added interests were the studies carried on at the University of Wisconsin• These studies included among others: anionic and cationic flotation, magnetic separation, roasting followed by magnetic separation, electrostatic separation, a hydrogen reduction process,6,7 and caustic digestion• The idea of removing silica by caustic extraction is not new•-1; The principle of concentrating siliceous iron ores by this method is based on the fact that silica is soluble in sodium hydroxide solution while iron oxides are essentially insoluble• For this process to be economically feasible, it is believed that some method of recovering silica from solution as a marketable by-product, with recycling of spent liquor is necessary. Precipitation of silica with carbon dioxide will allow subsequent causticizing of the spent liquor with lime• Caustic thus produced can be recycled for digestion• A flowsheet of the envisioned process is given in [Fig• 1] The purpose of this paper is to describe results obtained from the investigation of carbon dioxide precipitation of silica from sodium silicate solutions, which could be later applied to the caustic extraction process to treat low grade siliceous iron ores, particularly taconites.15 Theoretical Considerations The soluble silicates of sodium include sodium orthosilicate (2Na2O•SiO2), sodium metasilicate (Na20•SiO2), sodium disilicate (Na1O•2SiO2), and more highly siliceous products approaching Na2O•4SiO2• These materials have been described in detail by Vail•16 Stones has reported that the extraction of silica is not significantly affected by the amount of silica in solution until the Si02/Na20 ratio exceeds that of orthosilicate (Ratio = 1/2)• However, as the Si02/Na2O ratio exceeds that of metasilicate (Ratio = 1), silica extraction is significantly impaired• The present experimental work was confined to orthosilicate and metasilicate solutions, as to be successful the envisioned commercial process for the extraction of silica from siliceous iron ores would operate in this region. The mechanism of extraction of silica from aqueous sodium silicate solutions is complex in nature. For the carbon dioxide process-18 of core and mold making in which silica is also precipitated from sodium silicate solutions by gassing with carbon dioxide, the reaction of aqueous silicates with gaseous carbon dioxide has been represented by the equations: [ ] However, these equations are very much oversimplified• It has been shown that the reaction between aqueous sodium silicates and carbon dioxide gas is ionic in nature 18 Considering this fact and the mechanism for digestion of silica in aqueous sodium hydroxide, the overall conversion of aqueous sodium silicates to solid silica may be represented in ionic form by the following equations: [ ] Silicon in the silicate ion is shown here with a coordination number of 6• This may be in doubt, but it is known that ion (SiF°)-2 exists; it seems logical that the Si (OH) e 2 ion should be equally capable of existence•

Jan 1, 1977

By R. R. Bryan

SINCE the first applications of cyanidation to silver ores about 1906, treatment of ores in the Pachuca district has been entirely by straight cyanidation. Until about the year 1921, Real del Monte removed a sulphide concentrate (largely pyrite) by tabling in the cyanidation grinding circuit. The ores milled prior to the year 1935 were dry ores, ideal for cyanidation, and yielded 90 pct of the silver and 92 pct of the gold. Since 1934 there has been a notable increase in the refractivity to cyanidation of the average ores received by the mill. Tonnages of the former amenable ores have gradually been replaced by tonnages of more refractory ones, so that present average recoveries have fallen to 83 pct of the silver and 90 pct of the gold. During this period there has also been an increase in the content of base metal sulphides' in many of the ores although this increase is not always directly related to the increased refractivity. Since March 1950, the Cia. de Real del Monte has been developing a flotation treatment for some-of its ores before cyanidation. Suitability for flotation is judged not only by the base metal content, but also by the amount of additional recovery of silver and gold made possible by the combined treatment. Selection and segregation of ores most suitable for flotation before cyanidation is a difficult problem for the mines, and the best that can be done is a compromise that leaves much to be desired. Development of Pre-cyanidation Flotation at Real del Monte Much laboratory work has been done on flotation of ores from the Pachuca district, however, the data could not supply reliable information on the many factors connected with the economics of low-base metal-high silver ores. Pilot plant operations would be required. Pilot plant facilities were made available in March 1950, when it became necessary to close down Loreto's 100 ton per day selective flotation plant because of exhausted ore supply. The results of the 100-ton pilot plant tests indicated that a large part of our tonnage of cyanidation ores could be profitably treated by flotation, and a tonnage of 1100 tons per day was selected as the maximum tonnage of ore that the mines were likely to deliver, and because this tonnage fitted with the grinding capacity of this circuit. Equipment was ordered for the production of a single flotation concentrate. The conception of flotation before cyanidation, at that time, was the production of a single concentrate to be shipped to the lead smelter. The 100-ton pilot plant was immediately increased to 350 tons by installing additional flotation cells.. This. plant was operated continuously until the 1100ton plant was put in operation in January 1951. In July 1950, an increase in the quotation for zinc called for an investigation- of the possibilities of producing a zinc concentrate, and laboratory tests on the tailings from the 350-ton plant indicated that a

Jan 1, 1952

By E. W. Davis

About a year ago several carloads of magnetite ore from New York state were sent to the Mines Experiment Station at the University of Minnesota for grinding and concentration tests. The flowsheet used minus ½-in. ore as feed to a 3 by 3-ft. overflow cylindrical ball mill operating at 37 r.p.m. in closed ciruit with a vibrating screen. The screen oversize was returned to the mill for further grinding and the screen undersize was concentrated magnetically. Several different screens, between 8 and 35 mesh, were used and it was desired to determine the capacity of the ball mill operating in closed circuit with each screen. From earlier tests using this flowsheet and equipment, it was known that when comparatively coarse screens were used it was necessary to increase the grinding dilution in the mill beyond that ordinarily used when grinding in closed circuit with classifiers. It was also known that after the circulating load had passed about 100 per cent of the original feed it tended to increase rapidly in amount, and grinding conditions appeared to become unstable. These conditions may be peculiar to this small ball mill but they were very marked in this series of tests. If conditions were not right, the mill would suddenly "go dead," and it was found desirable to operate so that the occasional rattle of balls against the lining of the mill could be detected. As the test proceeded, it was found that the cure for almost any difficulty that developed in the grinding circuit was to add more water to the mill feed. To begin with, the mill was operated at pulp densities of from 60 to 65 per cent solids but gradually it was found that with higher dilutions grinding conditions were more stable and the ball-mill power per ton of screen undersize decreased. Actually, the most satisfactory operating conditions were secured with a pulp density of about 40 per cent solids. The ore was hard but it was possible to grind ½-in. feed so that all would pass an 8-mesh screen with a power expenditure in the ball mill of less than 2½ kw-hr. per ton and with a circulating load of less than 50 per cent. With so high a dilution in the feed to the mill, it was feared that steel wear would be very high. However, the mill was not noisy, and in order to find exactly the grinding conditions, a special test was made at the end of the main series of tests. After the mill had been in operation for several hours and grinding conditions had become balanced, the power to all the equipment was shut off at a given signal. The ball mill made about one revolution after the power was turned off, and just as soon as the mill stopped, a spout was placed at the discharge trunnion and all of the pulp that flowed out of the mill was caught in tanks. The manhole in the mill was then opened and the whole mill charge dumped into other tanks. Both of these products were weighed, dried, and weighed again. The balls were screened out and weighed, and the dry solids were sized. In this way it was possible to determine the nature of the pulp and the ore within the mill under actual grinding conditions.

Jan 1, 1947

By J. D. Edwards

When Charles Martin Hall invented the electrolytic process for the production of aluminum, one basic requirement was a supply of pure alumina. Now, more than 50 years later, the same requirement still holds, and there are no technical straws in the wind to indicate that this situation is likely to change. Long ago, Aluminum Company of America took a number of steps to see that this country should not face shortages of alumina and aluminum in time of emergency. As one step to this end, the Company developed and used a supply of bauxite in South America in order to conserve dwindling Arkansas reserves. Another, and more a program than a step, was to investigate every type of process that might bear the possibility of utilizing the low-grade ores of aluminum that are more plentiful than bauxite. It is a long story, but the investigation included electric-furnace processes, acid extraction processes and alkaline extraction processes, with or without preliminary furnace and sinter treatments. Some of the processes developed were operated in commercial production. This work has been continuous since the first World War. The turn of the wheel came sooner than was expected when, in 1942, bauxite shipmerits through the caribbean became uncertain and hazardous through the operation of submarines. The tremendous requirements of aluminum for the war effort were consuming high-grade bauxite at an alarming rate and a national search for substitutes began. High-grade bauxites, for technical and economic reasons, have been defined as those containing about 7 per cent or less of silica. While bauxite of this grade was limited in amount in the United States, there was in sight a substantially greater volume of low-grade, high-silica bauxite not amenable to economic treatment by the Bayer process, but richer in alumina than the clays or other available low-grade ores. It was at this point, and after several years of intensive research on this particular phase of the problem, that Aluminum Research Laboratories brought to completion the invention of one of its research staff, which made practical the utilization of large reserves of low-grade bauxite in America. Bayer Process The Bayer process has been almost universally used for the production of alumina from bauxite. This process, in outline, comprises the digestion of hely ground bauxite in a solution of caustic soda (NaOH) under pressure and at an elevated temperature. Alumina in the bauxite goes into solution as sodium alumi-"ate and the impurities—oxides of iron, titanium and silicon—remain undissolved. The undissolved residue, known as "red mud,'' from its color, is separated and run to waste. AluKlinum hydrate (A12O3.3H2O) is precipitated from the solution of sodium aluminate, and the liquor containing the caustic soda and some alumina is returned to the Process for digesting more bauxite. The aluminum hydrate, after washing, is calcined at a high temperature to produce

Jan 1, 1949

By L. W. Saperstein, D. Yeung, L. B. Phelps

Introduction The success of preventive methods for the mitigation of coal mine drainage formation will require a marriage of the minimization of initial coal mine drainage formation, the minimization of toxic fluid formation after toxic material disposal, the minimization of the mobilization and transpor¬tation of formed toxic fluids, and the maximization of the rate of diultion and attenuation of the formed toxic fluids. This paper contributes to the understanding of factors af¬fecting mobilization and dilution and attenuation of formed toxic fluids. When water containing a salt content different from in situ pore water in the soil is passed through a body of soil, the concentration of salts in the outflow will gradually change composition in a manner that depends on the kinds of hydraulic processes that occur in the soil. These processes are molecular diffusion, lateral dispersion, longitudinal dispersion, chemical reactions or exchange, physical adsorp¬tion, and biological retardation. This note considers only the first three processes. These three microscopic processes can be analyzed statistically on a probability basis. Such results, confirmed experimentally, indicate that a tracer introduced instantaneously as a point source will tend to form a Gaussian normal distribution as it moves downstream (Todd, 1964). Dilution During Flow The closed-form soltuion for pulse or instantaneous injec¬tion of a tracer during confined flow in a homogeneous porous medium shows that the maximum tracer concen¬tration at a point in space, x, after the time interval, t*, is given by [ ] where Aa is the effective flow cross-sectional area; A is the column cross-sectional area; a is soil porosity; M is mass of the injected tracer; C is concentration of the tracer in the permeating fluid and is a function of x and t*; D is the longitudinal dispersion coefficient; and va is the interstitial fluid velocity (Klotz and Moser, 1974). Therefore, the following conclusion can be made: • For a medium with a constant longitudinal dispersion coefficient, D, a constant soil porosity, a, a constant cross¬sectional area, A, and a fixed mass of injected tracer, M, Cmax will decrease as the time of peak outflow concen¬tration t* increases. • For a medium with a constant longitudinal dispersion coefficient, D, a constant soil porosity, a, a constant cross¬sectional area, A, a fixed mass of injected tracer, M, and a constant interstitial fluid velocity, va Cmax will decrease as the distance of travel, x, increases. • For a medium of constant cross sectional area, A, con¬stant soil porosity, a, a constant mass of injected tracer, M, and a fixed distance of travel, x, Cmax will decrease as D/va increases. Klotz and Moser (1974) concluded that D is about proportional to no.3 where r) is the kinematic viscosity; D is about proportional to a-3 where a is soil porosity; D in¬creases with increasing grain size; D increases as grains become less spherical; D increases with greater grain rough¬ness and angularity; D increases with increasing uniformity coefficient; and D is significantly affected by grain size and uniformity coefficient. The remaining parameters are not significant. Numerical Simulation Studies Numerical simulation studies were used to investigate the nature of flow and the optimum length of a toxic package within a saturated porous medium of fixed dimensions and configuration (Fig. 1). The width and position of the package was maintained constant (Phelps et al., 1981). Results indicate that if the package extends throughout the length of the model, the flow is vertically downward and

Jan 1, 1984

By C. H. Ma, J. S. Makris

This paper describes a modified X-ray diffraction microscope wzethod based on the general principles of Berg-Bawett's Method, Lang's Method, and Woosters1 Technique. This new method may be used for stationary or oscillatory photomacrographs, with both reflection and transmission techniques. In the stationary technique, a small point-source of a divergent X-ray beam of white radiation is used; in the oscillatory technique, a monochromatic divergent beam is employed. The chief advantages of the method described herein are that only approximate alignment of the Bragg angle -of-interest is needed during exposure, and that a large portion of a crystal-specimen surface can he irradiated. In recent years, the study of dislocations through X-ray diffraction microscope techniques has received considerable attention as concerns the per,fection of single crystals. Very promising results have appeared in the literature.'-' These techniques are very useful when the dislocation density of a crystal is near 105 per sq cm or lower. The principle of X-ray diffraction microscopy is based on the differential effect of X-ray primary extinction between the distorted parts and the perfect parts of a crystal. When an X-ray beam impinges on a set of planes of a crystal at the correct Bragg angle, diffraction of X-rays will occur with a 90-deg phase shift of the diffracted beam. This once-diffracted beam still forms the correct Bragg angle with the planes ahead of it, if the planes are not distorted. Thus, part of the strength of the once-diffracted beam is rediffracted and the beam continues to travel in the direction of the primary X-ray beam. This twice-diffracted beam not only takes away part of the strength of the once-diffracted beam, but also causes destructive interference with the primary beam because it now is exactly 180 deg out of phase with the primary beam. As the primary beam goes farther and farther into the crystal, its intensity is continuously weakened. Asa result, the ultimate X-ray diffraction power of a set of undistorted planes is largely extinguished. This phenomenon is called primary extinction.1,2'4 If the set of planes ahead of it is distorted, the once-diffracted beam will have less chance to be rediffracted, because it could not meet the exactly correct Bragg angle for rediffraction. If a part of the beam does rediffract, this twice-diffracted beam is not exactly 180 deg out of phase with the primary beam, and thus it causes less destructive interference. As a result, the registered X-ray diffraction power of the distorted part of a crystal appears stronger than that of the perfect part. Thus, point-by-point contrast of a diffraction image is theoretically possible from a crystal containing dislocations. Two most popular X-ray diffraction microscope techniques today are Berg-Barrett's Method and Lang's Method. Both of these techniques require a

Jan 1, 1964