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By Wilhelm Flesch

THE object of development work in the field of gasification is to find a process by which all' fuels, regardless of their individual properties, can be gasified economically. This paper describes the development of the Winkler process. The economy of the gasification process depends not only on the degree of the gasification effect but also, to a decisive extent, on the costs of fuel, production, and installations. Fig 1 shows the relation between fuel costs and gasification efficiency for various types of fuels. Since the fuel cost in the gasification of coke amounts to two thirds of the price of gas, and the thermal unit in brown coal costs only one third of what it costs in coke, it is easily seen how profitably cheap fuels can be gasified, especially if these fuels require minimum preparation and gasification is done in the vicinity of the mines. These possibilities were a great advantage of the Winkler generators, which were installed 20 years ago in the central German brown-coal fields and, at that time, were without rival for the gasification of finely divided fuels. In the Winkler generator (Fig 2), crushed and ground solid fuels were, for the first time, successfully gasified with oxygen and steam on an industrial scale. Unlike the usual processes formerly employed on an industrial scale, this gasification, which proceeds continuously in the glowing, seething fuel bed, makes only extremely low demands on the fuel. The fuel no longer has to be in lumps and abrasion-resistant, but must be capable of reaction and must be fed into the generator in a grain size of 0 to 8 mm. Brown coal or bituminous coal, or their low-temperature cokes, can be gasified in the Winkler generator. Large plants using fuels are operating on an industrial scale in Europe and in Japan. Table 1 shows the results obtained by gasifying these, fuels in Winkler generators. The thermal efficiency of gasification, which is the relation between the calorific value obtained in the gas produced and that of the fuel consumed, gives no indication of the quantity of steam generated in gasification. The generation of steam in the waste-heat boiler of the Winkler generator adequately-covers the steam consumed in gasification. To obtain effective gasification in the Winkler process, combustible mate-

Jan 1, 1953

By O. U. Bradley

The extent of losses, due to evaporation, sediment, and water, in crude oil stored in steel tanks, is a very interesting question, and particularly so at this time, when every reasonable measure should be employed to eliminate all possible losses of this important natural product. Available information on this subject is incomplete; e.g., during the development of the Cushing Field, considerable surplus oil was stored in steel tanks, but from time to time, owing to changes in weather, some losses by evaporation and short storage room on the leases, these tanks were topped out, thus rendering inaccurate any system designed to determine the average rate of evaporation or other losses of the oil over a given period. Furthermore, the losses, as shown by the records at the time, as a result of hasty gages, failed to take into account the temperature and gravity of the oil. Losses may be classified roughly as occurring from evaporation, presence of sediment and water, and leakage. The coefficient of expansion and contraction of crude oil, in relation to temperature conditions, is of material importance; also, the rate of evaporation is dependent upon the gravity of the oil, as the escape of the lighter hydrocarbons in fresh Cushing oil, of from 40 to 42 gravity, on a warm day is considerable, and from available records of temporary storage, it is safe to assume that a loss of 1 to 11/2 per cent. in volume will easily occur in light crude oil of Cushing grade over a period of 6 months, including the summer season. The presence of sediment, water, and other impurities, in the oil will also cause more or less deterioration in quality, particularly when the fresh production is run direct from the gage tanks, or from the wells on the lease to steel storage, as is often the case when an oil field is being rapidly developed. The bad oil will settle to the bottom of the tanks and may be determined by thiefing and running a centrifuge test on the sample. The amount of water in the oil depends entirely on the conditions under which it is produced. Sometimes in large producing wells, with heavy gas pressure, some water may come in with the oil, particularly if the well has been drilled to the top of a water sand. Under these conditions, cut oil is often produced, which, if not settled out, is carried over into the steel tanks. The losses by leakage in steel tanks are difficult to determine, for

Jan 1, 1920

By Theron Wasson

The state of Michigan showed considerable activity during 1935. The center of greatest interest was the Crystal field, which was unknown until March, 1935, and is an illustration of the effect of the unknown and unexpected on Michigan oil development. The fields of the state (Fig. 1) produced a total of 15,770,000 bbl. of oil in 1935, or about 5 million more than in 1934. In order to keep up this production 307 new oil wells were drilled, of which 88 were in the Crystal area, which produced 3,600,000 bbl. This field had a rapid development and reached its peak in October. Since that time increasing water encroachment has caused a rapid decline. The pay horizon at the top of the Monroe dolomite is only a few feet thick. Big producers were offset by dry holes or small wells. The discovery well, which was in section 11, Crystal township, Montcalm County, was drilled March 28, 1935. Its initial production was 5000 bbl. per day, and up to Dec. 31, 1935, had produced 350,000 bbl. of oil—dose to one-tenth of the total production of the field to that date. The Beaverton field, which was discovered at the close of 1934, did not develop rapidly. Four small producers were finished in the Dundee lime during the year. West Branch was active and at the end of the year had 97 producers, 72 of which were drilled during 1935. Most of the wells produce from the Dundee lime; a few are finished in the Traverse. The Birch Run field deserves special mention, for it is the one area where consistent production continues to be found in the Berea sand at about 1500 ft. Wells are small, 20 now producing a total of about 300 bbl. per day of oil of fine quality. The old Central Michigan producing area, extending from Nit,. Pleasant through Yost to Porter, continued to be the principal producing area of the state and accounted for 10,370,000 bbl. from 660 producing wells. Ninety-three new wells were drilled in this territory. The Porter field, at the southeast end of this producing trend, was outstanding from the standpoint of production. The old Leaton field of Isabella County added several new producers. Its production was greater than in 1934.

Jan 1, 1936

By P. D. I. Honeyman

CHANGING trends in concentration, involving regrinding with the subsequent production of a finely divided, high-grade flotation product, presents a real problem to the modern copper smelter. In the treatment of such a material, handling difficulties are numerous and potential possible losses may be very great. In many plants, of course, preliminary roasting of the furnace charge remains an economical necessity. Under such circumstances, the operator has no alternative except to follow orthodox roaster reverberatory practice, though taking all possible precautions to avoid excessive dust losses and bad working conditions. In other plants, roasting is not only unnecessary but would be highly undesirable, and at best the charge would be merely dried before entering the reverberatory furnace. In the latter case the possibilities of raw charging loom large. The Miami plant of the International Smelting Co. handles the entire output of concentrates of the several mining companies operating in the Globe-Miami district. All of these companies produce a flotation concentrate which, at the Miami Copper Co. particularly, is all reground and is high in grade. In addition, the Inspiration Copper Co. ships a substantial tonnage of high-grade cement copper to the smelter from its leaching plant. Following the flow sheet of the original plant, all such material received was bedded, dried in Wedge furnaces and smelted in reverberatory furnaces. At times during the history of the plant it has been necessary to carry the drying operation further, to a partial roast, before reverberatory smelting. Since the charge contained over 80 per cent flotation concentrates, the handling of which in a dried condition was at best a dusty job, unusual precautions were taken to cut down dust losses. The Wedge furnace plant was equipped with a Cottrell treater and all hoppers and calcine cars were carefully designed to cut down dust loss in transferring the so-called calcine from hoppers to car and from car to furnace. In the furnace itself the dusting problem was a severe one. Arch erosion was very rapid and flues and waste-heat boilers rapidly filled up with a partly smelted flue dust.

Jan 1, 1932

By William E. Bonini, Eugene A. Hickok

IN the course of an investigation directed toward expanding ground-water facilities in Essex and Morris counties, New Jersey, the Board of Water Commissioners of the city of East Orange authorized a seismic-refraction survey1 for the purpose of delineating bedrock topography below unconsolidated overburden. Results of the survey were highly satisfactory and led to the preparation of a comparatively detailed bedrock contour map. Knowledge of the bedrock depth and configuration was an important aid in selection of sites for test drilling. The portion of the East Orange Water Reserve under consideration is in the flood plain of the Passaic River about 10 miles west of Newark, N. J. The flood plain is about 175 ft above mean sea level and is bordered by low hills rising to elevations of approximately 250 ft. The bedrock underlying the Water Reserve consists of sandstone and shale of the Triassic Brunswick formation and is covered everywhere by deposits of unconsolidated glacial outwash sand and gravel, lacustrine clay, and recent river silt as much as 150 ft thick. Yield of wells in the sandstone and shale averages 100 to 200 gpm. Since production wells constructed in the sand and gravel aquifer in the buried river valley shown on the contour map (Fig. 1) yield 300 to 1400 gpm, it was proposed to locate additional production wells in this buried valley, where the yields per well would be maximum.

Jan 4, 1958

By R. V. Ramani, R. M. Mishra, Y. J. Wang

The validity of the concept of mine resistance in ventilation systems is examined. Mine resistance is the proportionality constant relating the head generated by a fan and the quantity flowing through it when the fan is connected to a mine circuit. It is the total effect of all branches in the mine ventilation circuit. Its importance lies in the fact that for new mine projections, and in the extensions of existing mines, ventilation calculations are usually based upon adjustments of the mine resistance. In ventilation circuits with regulation, and/or with multi-fans, simple adjustments of the mine resistance curve do not lead to correct solution. A ventilation system consists of several interconnected subsystems where one subsystem behavior affects that of the others, and the total system. Through a series of simple but realistic examples, the authors demonstrate the importance of the systems concept in mine ventilation, and its practical application to coal mine ventilation planning. Starting from an initial system with no leakage and one fan, more complex ventilation networks with several leakages, sections and fans, are analyzed to show that ventilation system design on the basis of total systems approach is the correct procedure.

Jan 1, 1979

By E. E. Carpenter

(San Francisco Meeting, September, 1915) BELIEVING that the results accomplished in the mill of the Churchill Milling Co., Wonder, Nev., during the fiscal year ended Sept. 30, 1914, will be of interest, I am presenting the more prominent facts and figures showing the progress made in treating this ore with especial reference to tonnage, supplies, costs, and extractions. The mill is situated at Wonder, Nev., in the Wonder mining district, 55 miles east from Fallon, Nev., the nearest railroad point. All supplies are hauled this distance by freight teams, this method of transportation having been found more reliable for continuous service than trucks, the price paid for hauling being slightly under one cent per pound. In this mill are treated the ores of the Nevada Wonder Mining Co. ,

Jan 6, 1915

By R. G. Knickerbocker

The copper smelter and refinery of The Messina (Transvaal) Development Co. Ltd., at Messina, South Africa, was erected in 1920 and 1921, but initial operations were deferred until late in 1922 on account of the collapse of the copper market in the interim. Because the anticipated production was to be only 20 tons of copper per day from a high-grade matte containing 60 per cent copper, and because the ore did not contain gold or silver in quantities sufficient to pay for separation from the copper bullion, it was decided to design the plant along lines quite different from the modern smelteries treating large quantities of sulfide ores carrying appreciable amounts of gold and silver. The plant was built, therefore, for the production of copper by the Nicholls-James process. By this method two-thirds of the copper matte produced was roasted in a Merton calciner and charged into a bath of the remaining one-third portion of unroasted matte in a reaction furnace. The original design was not only novel in providing for the Nicholls-James process, but also in comparison with new plants built elsewhere, because of factors inherent in the locality. Sources of supply for machinery were very remote; it takes three months after the placing of an order to accomplish delivery of machinery into Messina from the United States. Native labor was available at 30 to 40¢ per 8-hr. shift. These factors permitted plant design which provided no overhead cranes, no matte ladles, no slag pots or other usual mechanical means of transportation between departments. The design resulted in a compact simple plant with furnaces arranged close together and terraced so as to provide for gravity transfer of molten material from one furnace to another and for manual disposal of products. Fig. 1 shows the relative location of furnaces. The three furnaces are parallel on successively decreasing elevations. A superimposed feed floor was provided above the furnaces. All three furnaces were provided with coal-dust burners served from a Bonnot system of coal-dust transportation, and the waste gases were discharged into waste-heat boilers. A short trial of the Nicholls-James process developed several facts which brought about a decision to abandon the process. Owing to the

Jan 1, 1934

By Carl O. Thieme

INDUSTRIAL specialization has rapidly created a demand for new and better alloys. A more thorough understanding of the requirements of specific industries and the discovery of processes by which it has been made possible to utilize certain metallic elements in larger quantities, together with greater purity in most virgin metals, has resulted in a great increase in the number and variety of alloys. This increase in the number of elements entering into non-ferrous alloy manufacture has increased the contamination in secondary metals, as received by the smelter. In order to successfully refine scrap of a much more complex nature it has been necessary for the smelter to install new equipment and to constantly search for new methods of refining as well as to improve the methods already in use. In some instances it is largely responsible for specialization in the secondary metal industry. It is usually necessary to remove the contaminating impurities in scrap or eliminate to a degree not objectionable, no matter how introduced. In order to know what metals or elements are of an objectionable nature, the smelter must make a study of the effect of single elements and combinations of elements on many of the physical and mechanical properties, such as strength, ductility, color, hardness, casting soundness, etc. The smelter of secondary metals must be able to utilize scrap materials of all kinds, for, sooner or later, most fabricated metal parts and the waste products formed during the process of manufacture find their way to the smelter. In addition, the smelter, in contracting for salvaged material from the producer, gets some undesirable and unusual materials. The value of secondary metals and the importance of their recovery has increased in recent years and it is no longer sufficient to recover the base metal only, such as copper from copper-bearing material, but it is essential that other elements present be recovered and converted into useful form. This paper will deal not with the smelting and refining or separation of impurities from secondary metals but an attempt will be made to show the increased value of metal scrap when properly segregated and in a less contaminated condition. There are five principal reasons for contamination in secondary metals: New and unusual alloys, productive methods,

Jan 1, 1928

By Harry A. Lipsitt, Rollins E. Hook, Attwell M. Adair

A necessary condition for a uniaxial stress distribution in compression testing is that the specimen end surfaces make full contact with the apparatus compression plates. In addition all compression surfaces must be machined perpendicularly to the loading axis. In general this condition is not sufficiently fulfilled in a simple apparatus. Therefore on application of the load the compression plate may first touch the specimen at one point only. This may cause crumbling of the specimen if the material under investigation is brittle (e.g., ionic crystals). It also precludes a precise study of the beginning of deformation as the effective cross sectional area is unknown. In addition a complex stress state is set up which renders quantitative work impossible, e.g., on dislocation density and arrangement as a function of stress. To overcome these difficulties a "self-adjusting" upper compression plate is required. A simple and reliable method of achieving this is shown in Fig. 1: A steel hemisphere is placed on the specimen. As soon as the compression plate (which no longer needs precise machining) makes contact with this hemisphere the load is uniformly distributed over the whole specimen surface, and a homogeneous uniaxial state of stress results. An advantage of the method is that the loading is always central, even when the specimen surface is not parallel to the compression plate (Fig. 1, dashed line). Therefore to a first approximation a uniaxial state of stress is maintained if the misorientation of the specimen end surface is small. Two typical stress strain curves of identically prepared sodium chloride crystals are shown in Fig. 2; curve 1 was taken without and curve II with the hemisphere under otherwise identical conditions. The improvement in the accuracy of defining the beginning of elastic and plastic deformation is clearly indicated. The writer wishes to thank the Deutsche Forschungsgemeinschaft for financial support, and Professor P. Haasen for providing facilities for this work.

Jan 1, 1962

By P. C. Claspy, P. H. Smakula

The electro-optic coefficients of LiNbO, and KTao.,5Nboe3,O3 (KTN) have been experimentally determined at the HeNe laser wavelengths of 0.6328, 1.15, and 3.39 µ. The coefficients were calculated fronz the retardation induced by an applied electric field. The retardations were measured by observing one or move transmission maxima and minima as a function of the applied field, and by measuring the retardation directly with a Babinet-Soleil compensator. The values obtained agree quite well with those previo~lsly reported by others. Extinction ratios as high as 3000:l have been observed. Resistivities and Curie temperatures have also been measured. Light-induced optical inhornogeneities in LiNbO, repovted by others have been observed. CONSIDERABLE interest has recently been shown in the electro-optic properties of LiNb03 (Refs. 1 and 2) and KTN.3, 4 These crystals have a number of advantages over such well-known materials as KDP. These advantages include: 1) the infrared transmission of KTN and LiNb03 extends to 5 p whereas that of KDP cuts off at 1.5 p ; 2) driving voltages required for modulators using these materials are considerably lower than those required for KDP. After first discussing some theoretical considerations of the electro-optic effect, we will report the results of measurements of the dc electro-optic coefficients of these materials at wavelengths of 0.6328, 1.15, and 3.39 p. THEORY The generalized equation for the electro-optic effect can be written: Although lattice polarization is the fundamental variable, rather than the applied electric field, we follow the convention of expressing the linear electro-optic effect in terms of the applied field. For LiNbO the quadratic terms are negligible. Since KTN has a center of symmetry the rijk tensor is 0. After applying the usual symmetry considerations and index contractions,* the rijk and gijkl tensors can be expressed as follows: LiNbOaT If field directions are now chosen, the change in refractive index and the angle between the induced optic axes and the crystallographic axes can be found. The induced rotations of the principal axes of the index ellipsoid for fields in principal crystallographic directions are given in Table I. Since KTN is initially isotropic, the principal axes in the zero-field condition are parallel to the crystallographic axes. The induced retardations are given in Table 11. EXPERIMENTAL We obtained the various electro-optic coefficients by measuring the induced retardation as a function of

Jan 1, 1968

By J. L. Bruce

For a number of years, from an economic standpoint, the ventilation of the Colorado mine of the Davis-Daly Copper Co. has been a difficult problem. The development through the main hoisting shaft was pursued for a number of years and to a depth of about 2500 ft.( 762 m.)before large bodies of good commercial ore were opened up. This lack of early success in exploration restricted expenditures for ventilation, with the result that the present orebodies were opened up at a depth of 2500 ft., and in country rock f relatively high temperature, without adequate means f supplying the development faces and the stopes with cool or fresh air of low humidity and suitable velocity. During the progress of development, two or three connections were made at higher levels with the workings of adjoining operations, but these were of a makeshift and unsatisfactory character and afforded no means for carrying downcast air from the surface, in anything approaching a direct course, to the working faces. During the latter part of 1917, arrangements were made with. the Anaconda Copper Mining Co., whereby, during the year 1918 and the early part f the year 1919, a crosscut was driven eastwards on the 2500-ft. level of the Colorado shaft and connection was made by means f an up-raise of approximately 200 ft. to a crosscut driven westwards from the 2800-ft. level of the Belmont mine, this crosscut at the top of the raise being driven at Davis-Daly expense by the Anaconda Copper Mining Co. This connection, in conjunction with a No. 11 Sirocco fan installed near the Colorado shaft, provided the Colorado mine with a maximum of approximately 30,000 cu. ft. per min. of relatively fresh air, most of which came in a fairly direct route from the downcast Belmont shaft through the No. 11 fan to the 2500-ft. level stopes. After passing through these stopes and the stopes of overlying levels, it passed through the main crosscuts on higher levels back to the Colorado shaft, which provided the upcast for this current of air. The improvement

Jan 9, 1921

Located 19 km (12 miles) north of the Carlin gold mine and 43 km (27 miles) north of Carlin, Nevada, is Carlin Gold Mining Company's Bootstrap mine. This deposit is similar in character to the Carlin ore, except it occurs in the upper plate formation of the Roberts thrust fault rather than in the lower plate rocks. The deposit was discovered over 50 years ago and operated on a small scale in the early 1960s using vat cyanide leaching. Newmont Mining Corporation acquired the property in 1967, and from an open pit mined 650,000 mt (717,000 st) of ore which was trucked to the Carlin conventional cyanide plant. (See Gold and Silver Cyanidation Plant Practice by F. W. McQuiston, Jr. and R. S. Shoemaker, Society of Mining Engineers, AIME, 1975. ) With the same grade of ore, the gold recovery by conventional methods is somewhat lower than that of the Carlin ore, even with a finer grind. Leaching tests were first conducted by the United States Bureau of Mines in 1968; and in 1971, ‘72, Carlin conducted extensive cyanidation investigations by bottle testing, simulated heap leaching in tall columns, and small scale pad heap leaches. Gold extractions ranged from 55% to 95% with low reagent consumptions. In 1974, a 545 mt (600 st) heap leach test was made at Carlin on pit-run low-grade material assaying 1.37 g per mt (0.04 oz per st). In a six week period the extraction of gold was 47.57'0. Carlin uses atomic absorption analyses on cyanide leach liquors to determine potential leach ore grade rather than fire assaying because it gives a more practical assessment of leach recovery than does fire assaying, which accounts for the total gold content. Before open-pit mining was started, plans were made to heap leach, the ore at Bootstrap which was too low grade for transport to Carlin. The topsoil outside of the designed pit area overlaid ancient lakebed clays which were found suitable for compaction to form an impervious base on which the dumps to be leached could be placed. By using the cost of truck haulage and milling at the Carlin plant and the price of gold 'at that time, a cutoff grade was established which resulted in the gold content of the first 150,000 mt (165,000 st) of the heap leaching ore being 1.51 g per mt (0.044 oz per st). An additional 500,000 mt (552,000 st) of lower grade material, assaying 0.96 g per mt (0.028 oz per st), is available for leaching as economics permit.

Jan 1, 1981

By D. A Dahlstrom

THE necessity for slime elimination from valuable mineral and coal products has become increasingly significant within the past 5 years.' , Most of the mechanized mining and present beneficiation methods have resulted in larger tonnages of extreme fines, silts, and clay, commonly called slimes, processed by the preparation plant. Economics in turn have forced the plant operator constantly to lower the bottom size of effective beneficiation, necessitating an almost complete recovery of valuable solids regardless of particle size. When these important factors are considered, the need for an efficient slime removal method for fine solids becomes self-evident. Ordinarily slimes have been classified as ranging from 75 microns, or 200 mesh, down to colloidal dimensions of less than one micron. At the same time there is general agreement that the most troublesome slimes are primarily clays, silts and extreme fines, so that the actual top size may be placed at approximately 30 to 40 microns. This immediately restricts any desliming operation to classification, as screening or other methods would be both uneconomical and grossly inefficient at such dimensions. The liquid-solid cyclone has been recognized as one of the most efficient fine size classifiers available to the mining industry. Applications of cyclones to thickening, high efficiency solid recovery, degritting,

Jan 1, 1953

By O. A. Knight

THE suppression of the austenite-martensite transformation that can be brought about by the addition of certain alloying elements, such as manganese or nickel, to plain carbon steel has been known for a long time, especially since Guillet's investigation at the beginning of this century. The extent to which the critical range is lowered depends upon the percentage and nature of the special element added. First the Ar1 point is lowered gradually, then at a certain percentage a sudden lowering to about 300° C. occurs. The addition of still larger percentages of the alloy may result in lowering the Ar1 point to room temperature or below, in which case the steel remains austenitic at room temperature, especially after rapid cooling. The effect of alloying elements of this general type on the lowering of the transformation temperature is well illustrated by Sauveur(1)§ in his Fig. 281a, reproduced here as Fig. 1. This type of curve is characteristic of both manganese and nickel though it requires about 2 1/2 times as much nickel as of manganese to produce the same result. The influence of nickel in retarding the transformation of austenite into the lower transition constituents is due to its lowering of the critical range of the steel, which is also well illustrated by Sauveur(1) (after Osmond) in his Fig. 284, reproduced here as Fig. 2. Scott(15) presents a diagram illustrating the effect of manganese on the Ac3 and Ar3 points which is strikingly similar to the diagram above referred to in Fig. 2 concerning the effect of nickel.

Jan 1, 1934

By W. L. Austin

THE advance made in recent times in this branch of metallurgy is indicated y the attention the subject is receiving from important American copper-producing companies. Reference to the files of publications devoted to the mining industry discloses that some 20 American companies are actively investigating the amenability of their ore, or other material, to leaching methods, and that plants of varying capacities up, to 10,000 tons per day are under construction or are projected. Several leaching works are in commercial operation. It has been frequently pointed out that no method of leaching is universally applicable, for the reason that each ore differs in some particular from apparently similar ore found elsewhere, and also because local conditions are rarely the same even where the ore closely approximates in character that being worked at some other point. For this reason the older companies are approaching the matter in a conservative manner, experimentally determining for themselves the salient features in each instance. In considering a leaching proposition three factors at once fix the attention: (1) disengaging the metal from mineralized forms in Which it is found in nature; (2) recovering the metal in a commercial state after it has been dissolved; (3) the apparatus best adapted to carrying out the several operations.

Jan 8, 1914

By Luis N. Henriquez

One form of encouragement for the small copper mining sector in Chile has been the construction and management of custom mill facilities to process the ore. The paper describes an economic analysis framework for the evaluation of government custom milling opportunities. Feasible mill locations and capacities within a region are selected and the infrastructure and milling costs associated with each combination of location and capacity are assessed. For each combination, the economics of developing and operating known copper deposits in the region is evaluated, and an optimum cut-off grade and mine capacity is selected for the deposits which justify development. Since the total ore supply from mines in the region may exceed a particular custom mill capacity, mine development is scheduled to optimize the economics of the mill facility. Finally, the best custom mill location and capacity is selected on the basis of overall costs and benefits. A copper mining region in northern Chile is used to illustrate the application of the model. A mobile mill approach is examined and compared with the fixed location mill alternatives. Results are analyzed as a function of copper price.

Jan 1, 1977

By D. A. Dahlstrom

THE necessity for slime elimination from valuable mineral and coal products has become increasingly significant within the past 5 years.' , Most of the mechanized mining and present beneficiation methods have resulted in larger tonnages of extreme fines, silts, and clay, commonly called slimes, processed by the preparation plant. Economics in turn have forced the plant operator constantly to lower the bottom size of effective beneficiation, necessitating an almost complete recovery of valuable solids regardless of particle size. When these important factors are considered, the need for an efficient slime removal method for fine solids becomes self-evident. Ordinarily slimes have been classified as ranging from 75 microns, or 200 mesh, down to colloidal dimensions of less than one micron. At the same time there is general agreement that the most troublesome slimes are primarily clays, silts and extreme fines, so that the actual top size may be placed at approximately 30 to 40 microns. This immediately restricts any desliming operation to classification, as screening or other methods would be both uneconomical and grossly inefficient at such dimensions. The liquid-solid cyclone has been recognized as one of the most efficient fine size classifiers available to the mining industry. Applications of cyclones to thickening, high efficiency solid recovery, degritting,

Jan 1, 1953

By Porter W. Shimer

Ox completing the filtration of a solution of pig-iron in hydrochloric acid, and after apparently everything had been transferred to the paper, a minute residue was observed remaining in the angle of the beaker in which the solution had been made. This residue had a steel-gray color and metallic luster and was hard and gritty when touched with a glass rod. Placed under a microscope and magnified 400 diameters, the substance was found to be made up almost entirely of opaque cubical crystals with some fragments having the same color and luster as the cubes. The appearance was strikingly like that of a collection of crystals of iron pyrites. The amount of material obtained from ten to twenty grams of borings by solution in dilute hydrochloric acid, and careful decantation from graphite, was barely sufficient for a microscopic slide. A qualitative test of the minute amount of material obtained in this way, showed the presence of considerable titanium. After many fruitless trials by decantation and other methods, sufficient material for quantitative analysis was finally obtained in the following manner: Two hundred and fifty grams of the coarsest possible borings were dissolved in seven liters of hydrochloric acid of sp. gr. 1.03. A solution by the aid of heat took four to five hours. It was allowed to settle, and the clear solution poured off. Great care was taken that none of the material which had settled to the bottom was lost. The graphitic residue was then completely transferred upon fine bolting-cloth, previously Stretched and fastened over the top of a beaker. All graphitic lumps were broken up by the finger while

Jan 1, 1887

By William Catron

BECAUSE the rich La Colorada orebody of the Cananea Consolidated Copper Co. does not outcrop at surface, after its discovery (by churn drill) and before mining was begun, a large amount of development work was done to learn the shape and extent of the ore and to provide for eco-nomic and effective mining. As these openings progressed, the actual shape of the orebody became known and with this knowledge came the realization that mining this rich and unusual deposit without loss of ore or danger to workmen presented several difficult problems. This paper will describe one of these problems; namely, the handling of an inverted cone-shaped core of barren rock, on the periphery of and below which the ore occurred. It was clear that as the ore was mined out from around and under this inverted cone of rock, weighing about 3,000,000 tons, it would hang over the workings, held suspended only by its own tensile strength and threatening to crash into the stopes below with disastrous results to men and mine. The method of mining developed during the years 1928, 1929 and 1930 anticipated and provided for the eventual collapse of this core of rock. In November 1936 the core settled clown gradually on the edges of the stopes, descending a total of about 50 ft. vertically. The men in the stopes had ample to warning and no lives were endangered. There were no air blasts in the mine. The core broke loose from its top support at the 500-ft. level along old fracture planes and evidently is crushed to the ninth level. Movement was of short duration; the following shift went to work in the stopes as usual. No ore was lost.

Jan 1, 1938