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By B. L. Sackett

IN a general way, the situation in copper smelting is quite similar to that of lead smelting. In other words, the introduction and successful flotation of copper ores has tended to decrease greatly the tonnage to be handled by the smelter and therefore has resulted in curtailed smelter operation. However, copper production has increased rather than decreased. The general effect of flotation has been to decrease the number of furnaces operating, but it has also increased to a considerable extent the difficulties in handling the material, just as was outlined in the case of lead smelting. In copper smelting, the flotation concentrates must be roasted and prepared for the reverberatory furnace. HANDLING FLOTATION CONCENTRATES The handling of copper flotation concentrate, which is more or less wet and mucky, is more difficult than was the handling of the old gravity concentrate. Also, the roasting of the concentrate and the handling of the calcine produced is more difficult. That is mainly because the material is extremely fine and the tendency to higher dust losses is greatly increased. The draft in the roasting furnaces must be kept as low as, possible, and the feeding of the calcine into the reverberatory furnace must be done carefully, so that dust losses may be kept at a minimum. TRANSITION STAGES OF CONCENTRATION AND SMELTING The general process through which the copper smelter has gone as a result of flotation has been somewhat as follows: In the relatively old days, when copper ores of low grade were first treated, a large number of furnaces were required, and naturally comparatively poor recoveries were made. Also, the profits on low-grade ores were small and the handling of those ores was possible only because of the large tonnages. Then came gravity concentration, which was a great improvement but which only incompletely removed the gangue material, the insoluble and the iron, from the copper ore, and made comparatively poor copper recoveries. This still left a considerable amount of insoluble and iron, as well as the copper itself, to be treated at the smelter.

Jan 1, 1928

By W. D. Robertson

THE effect of a tensile stress in accelerating the corrosion-cracking of certain alloys of aluminum, magnesium and iron is widely recognized. The literature is extensive and it is only necessary to cite the recent ASTM-AIME Symposium.1 There are, however, no experimental data conclusively showing the effect of an applied compressive stress. It has been inferred, from the generally accepted mechanism postulated for the acceleration of corrosion by an applied tensile stress, that a compressive stress would inhibit corrosion. Furthermore, it is observed that corrosion cracks invariably start on the tension side of a specimen bent as a simple beam. This observation has led to the suggestion that shot-peening of the surface would inhibit or eliminate stress-corrosion cracking in magnesium alloys, and aluminum alloys.3 However, the degree of inhibition produced by shot-peening does not appear in the literature except for the above qualitative references which are unsupported by experimental data. Moreover, while peening undoubtedly imposes a surface compressive stress, the alteration of the surface by the severe cold-working involved complicates the interpretation of the results with respect to the effect of a compressive stress alone. If corrosion proceeds along preferential paths produced by precipitation from solid solution, it is easily understood how a component of tensile stress normal to these paths will accelerate the corrosion and lead to ultimate failure. On the other hand, if the metallurgical conditions are such as to create preferential corrosion paths, then it is difficult to understand how a compressive stress, produced by shot-peening or any other method, can inhibit corrosion except in the sense that the accelerating factor is removed. Thus, if the basic corrosion mechanism remains unaltered, there should be little or no inhibition resulting from the application of a compressive stress if in¬hibition is defined as the reduction of the degree of corrosion below that experienced in the unstressed state. The experiment described below was performed to isolate and evaluate the effect of a compressive stress on the corrosion and stress-corrosion cracking of a high strength aluminum alloy (24-S); and, by a comparison of the corrosion rate of specimens subjected to an applied tensile stress, compressive stress and in the unstressed state, to clarify the role of stress in intergranular corrosion. EXPERIMENTAL PROCEDURE The experiment was conducted with the aluminum-copper-magnesium alloy known as Duralumin 24-S.* Specimens, 0.064 in. thick, in the form of I in. wide strips sheared parallel to the rolling direction,

Jan 1, 1947

By Nathaniel Herz

(San Francisco Meeting, September, 1915) THE use of zinc dust for precipitating the precious metals from cyanide solutions is well established now in many places, and has many advantages over the shavings method of precipitation. Although much work has been clone on methods of testing zinc dust, as yet no absolutely positive test has been devised to determine the value of the material. Certain facts are known, but often it happens that results on a working scale are not as expected from the laboratory tests. The purpose of this paper is to record some observations made while testing. a large variety of zinc-dust samples in the laboratory. W. J. Sharwood1 has described standard analytical methods, which have been followed, with the changes noted. The regular determinations include zinc oxide, by solution in ammonia and ammonium chloride, lead, and precipitating efficiency. In some cases cadmium was determined by a method similar to that described by J. E. Clennell.2 I will discuss the precipitating-efficiency test in detail. The procedure is as follows: "A solution of potassium silver cyanide is prepared by dissolving 10 grams of silver cyanide (AgCN) and 5 grams of `99 0' potassium cyanide in a little water and diluting up to 1000 cc. It is then adjusted by addition of a little more KCN or AgCN until the solution indicates from 0.12% to 0.15% free KCN by titration with standard silver nitrate. The titration is best made by using a 10-cc. or 20-cc. sample, adding 1 cc. of 2 % potassium iodide, and a slight excess of ammonia, as the end point is then sharper. Or 15 grains of pure crystallized KAg(CN)2 may he dissolved in a liter of water and 1.5 grams KCN added. "Weigh out 0.5 grain zinc dust into a 300-cc. beaker. Add a few cubic centimeters of water and stir until, zinc is well mixed, then pour in 250 cc. of the prepared solution, stirring vigorously. See that all lumps are broken up, and continue stirring for fully five minutes. Stir occasionally (at least every 10 minutes) until the end of two hours from the addition of the solution. Then filter upon an 11-cm. filter, wash precipitate thoroughly, sprinkle with test lead, wrap it carefully in the paper, plaice in a scorifier with about 20 grams test lead, burn paper cautiously in muffle, scorify five minutes, cupel at low temperature and weigh silver. Milligrams silver obtained from 0.5 gram zinc X 0.0606 = percentage precipitating efficiency."'

Jan 8, 1915

By Nathaniel Herz

The use of zinc dust for precipitating the pecious metals from cyanide solutions is well established now in many places, and has many advantages over the shavings method of precipitation. Although much work has been done o,n methods of testing zinc dust, as yet. no absolutely positive test has been devised to determine the value of the material. Certain facts are known, but often it happens that results on a working scale are not as expected from the laboratory tests. The purpose of this paper is to record some obscrvations made while testing a large variety of zinc-dust samples in the laboratory. W. J. Sharwoodl has described standard analytical methods, which have been followed, with the changes noted. The regular determinations include zinc oxide, by solution in ammonia and ammonium chloride, lead, and precipitating efficiency. In some cases cadmium was determined by a method similar to that described by J E Clennell.2 I will discuss the precipitating-efficiency test in detail. The procedure is as follows: "A solution of potassium silver cyanide is prepared by dissolving 10 grams of silver cyanide (AgCN) and 5 grams of 99 %' potassium cyanide in a little water and diluting up to 1000 cc. It is then adjusted by addition of a little more XCN or AgCN until the solution indicates from 0.12% to 0.15% free KCN by titration with standard silver nitrate. The titration is best made by using a 10-cc. or 20-cc. sample, adding 1 CC. of 2 % potassium iodide, and a slight excess of ammonia, as the end point is then sharper. Or 15 grams of pure crystallized XAg(CN)2 may be dissolved in a liter of water and 1.5 grams XCN added. " Weigh out 0.5 gram zinc dust into a 300-cc. beaker. Add a few cubic centimeters of watcr and stir until zinc is well mixed, then pour in 250 cc. of the prepared solution, stirring vigorously. See that all lumps are broken up, and continue stirring for fully five minutes. Stir occasionally (at least every 10 minutes) until the end of two hours from the addition of the solution. Then filter upon an 11-cm. filter, wash precipitate thoroughly, sprinkle with test lead, wrap it carefully in the paper, place in a scarifier with about 20 grams test lead, burn paper cautiously in muffle, scorify five minutes, cupel at low temperature and weigh silver. Milligrams silver obtained from 0.5 gram zinc X 0.0606 = percentage precipitating efflciency."³

Jan 1, 1916

By Charles White Merrill

A forecast of mineral demand during the remaining years of the 20th century can serve as an excellent starting point for student mining and mineral engineers in planning their professional careers. This approach of appraising demand before seeking supply is not new. of course, but, it is quite the opposite of the traditional attack on mineral resources in which adventurous prospectors, followed by practical miners and engineers, have in the past sought out and developed mineral supply with a faith that markets existed for the products they were preparing to offer. Such faith, even to this day remains fully justified if the product is gold, which still enjoys an unlimited market at $35 a fine ounce, but other mineral products must compete in markets of varying stability and capacity for adsorbing new supply. The world is in a period of rapid expansion. Most basic is the worldwide increase in the number of people which has reached proportions popularly termed "the population explosion." Conservative demographers, noting an increase from 1 ?h billion people at the beginning of the 20th century to 3 billion today, are predicting 5 billion by the year 2000; some predictions are much higher.

Jan 2, 1964

By John G. Roscoe

Continental Materials Corp. operates two under- ground uranium mines in the Crooks Gap district of central Wyoming. Both mines are developed as slusher mining operations through vertical shafts with single level haulage to slusher pocket loading stations. In general, operations at both mines are similar except for depth and some minor details. The shafts are identical in cross-section, each with a 4 ½ x 5-ft hoisting compartment and a 4 ½ x 5-ft manway. Kimberley-type skips are loaded from measuring pockets through air-operated guillotine gates. The Seismic shaft is 245 ft deep and the Reserve shaft depth is 585 ft. The Seismic shaft was sunk in 1957 and the mine has been producing 200 tpd since 1958. The Reserve shaft was sunk in 1960 and the mine is still in the development stage. The hoists of both mines are equipped for semi-automatic operation; that is, they are mechanically controlled to perform a certain cycle of duty and then stop until another cycle is manually initiated. The Seismic hoist, however, was originally a conventional, manually controlled unit; the reserve hoist was installed as an automated system.

Jan 7, 1962

By P. Barnes

These furnaces are of the ordinary Siemens type, and present no special peculiarities of construction. The bed of each is 8 feet by 20 feet clear inside of the walls and ports. The producers are placed at a distance of about 200 feet from the furnaces, and the gas is collected in an iron tube and led across the yard overhead. It then drops underground into the gas flue, and is distributed to the furnaces. A considerable weight of floor-plates over the valve-pits is included in account 39, but none of the general stock of floor-plates for the mill were charged to the furnaces. In Table No. 1 is shown the money cost of the furnaces as distributed to the several accounts named. In Table 2 is shown the proportion of each account due to each of the several items or classes of expenditure named.. It is thus rendered possible, almost at a glance, to determine the money value in this particular case of each of the items named. The regular work of three of these furnaces in heating steel blooms for the months of January and February, 1877, was 77 rounds per week of 60 blooms each, or 4620 30-foot rails per week. Each furnace will heat 8 14-inch ingots, for three rails each, at one time.

Jan 1, 1879

By F. H. Buttner, E. R. Funk, H. Udin

Gold wires, 5 mil in diam, are found to creep viscously up to approximately 5.5x106 dynes per sq cm around 1300°K. Beyond this point, an additional slip mechanism appears. The average coefficient in the viscous range at these temperatures is 18.9 ± 7x10 12 poises. This observation and other visual observations made during the experiments agree with Herring's predictions. EARLY creep studies by Chalmers on single crystals of tin produced some of the first experimental evidence to suggest that metals deform viscously under very light stresses. That is to say, over a small initial stress range up to 150 g per sq mm, the single crystals of tin elongated at a rate proportional to the stress. The proportionality constant between strain rate and stress is by definition the coefficient of viscosity. Chalmers named this range of proportionality the microcreep range. One of the first theoretical attempts to explain the microcreep phenomenon was given by Kauzman,' who applied Eyring'sQ eneral statistical mechanical theory of shear rates to creep data then available. From the rate theory, he deduced the general hyperbolic sine law for shear rate on the assumption that the distortion of the reaction path due to stress is negligible and that the system is an infinite continuum. In order to fit the available data to the equation, it was necessary to assume that creep occurs by shear of rather large blocks of material, rather than by an atomistic shear process as in diffusion. The size of the shear units was shown to increase with increasing temperatures and lower stress, or lower strain rates. In applying the hyperbolic sine law equation to the microcreep data of tin, it was shown by Kauzman that the unit shear blocks became quite small, which suggested to him that microcreep might be a self-diffusion process. Making this an assumption, Kauz- man reworked the rate theory and arrived at an equation where the shear rate is proportional to the stress. In the form of viscosity coefficient, the equation is: o- _ \LkT V = 2qAlA where u is the stress; S, the shear rate; X, the distance in shear direction moved by flow units relative to one another in the unit process; L, the distance between layers of units of flow; T, the absolute temperature; k, Boltzmann's constant; qAL, a constant proportional to the size of flow units; and kT A, the diffusion coefficient = ------kTe-Q/RT = K X2 h Kauzman applied the expression to molten lead with a fair degree of success, the calculated viscosity being low by a factor of four. For solid lead the result is 5000 times too high. Taking q = 1, and X = dA = L = 21 = nearest neighbor distance, Udin, Shaler and Wulff8 found the equation low by a factor of lo7 for solid copper. On the basis of the discrepancy in lead, Kauzman indicated that microcreep is probably not a simple diffusion process of molecular shear units, but one involving preferential diffusion paths, or even that diffusion in solids does not involve shear at all. Cottrell and Jaswon4 treat the microcreep phenomenon on the basis of slow moving dislocations and apply it to Chalmers' data on tin with fair success. Nabarro Suggested a microcreep model in which mass movement in the direction of pull is effected by the diffusion of holes. The tension on the crystal effectively creates a greater hole concentration at regions near the crystal surfaces perpendicular to

Jan 1, 1953

C. W. WASHBURNE, New York, N. Y. (communication to the Secretary *).-We are indebted to Mr. Pack for his detailed description of a rapid method of estimating the oil reserves in large fields. Although his method will give only a rough approximation of the reserves, nevertheless it is much more reliable than the saturation method. The saturation method should not be used when any other method is available. The paper describes a generalized method, which is especially useful in estimating the total resources of a Nation or State. Properly it is a method of estimating the oil that can be extracted by common methods within a given period of time. It does not account for

Jan 10, 1917

By T. H. Alden

Using alternating tension-compression straining, the hardening of magnesium oxide single crystals was studied up to large stresses and strains. At 0.25 pct plastic strain amplitude, the hardening curve is approximately linear with slope 25,000 psi from the shear yield stress, 7 to 8000 psi, to 35,000psi. Above this stress, the slope decreases. The strain hardening behavior of MgO is considered qualitatively similar to that of metal single crystals. The relatively high stress attainable by strain hardening is associated apparently with the high yield stress on the cross-slip system, (001) <110>. Cleavage fracture during testing is uncommon. It is argued that the centers of high internal stress at glide band intersections, at which cracks tend to nucleate, are dispersed by cyclic strain. Special features of the glide band structure produced by cyclic strain and revealed by dislocation etch pits, support this view. Strain hardened MgO has mechanical properties greatly superior to the as-received material: yield stress, greater than 100,000 psi; elongation to fracture about 1 pct. A material is said to strain harden if the yield stress increases with an increment of plastic strain. This definition is usually applied for straining done in one direction, but is also applicable when the strain direction is periodically reversed, Fig. 1. For certain metal single crystals, data are available which permit a comparison of the hardening behavior for cyclic straining and for tension straining.&apos;-4 With certain qualifications, these data show that the same processes of hardening are operative in each type of test.5 Despite this fact, the importance of the technique is not immediately evident, although tension-compression studies of the common metals appear to suggest some deficiencies in theories of strain hardening developed exclusively on the basis of tensile tests. However, a recent observation suggests that the cyclic straining method may be very useful for studying semibrittle crystals in which large plastic strains are not accessible in unidirectional testing. The observation is that zinc crystals, when strained in tension-compression at -52°C, do not fail by cleavage at low stress (-500 psi)6 as they do in tension, but harden to a limiting stress of more than 5000 psi over a total plastic strain of about 600 pct.2 An important characteristic of the behavior of zinc crystals is the high stress, relative to the yield stress, attainable by strain hardening. By comparison, the hardening of aluminum single crystals tested by an identical technique saturates at 1100 psi. This difference is best explained by the cross-slip hypothesis of dynamic recovery.7,8 In zinc, cross slip is difficult because of the high yield stress for glide on planes other than the basal plane in the < 1120 > zone. The present work was undertaken in order to test whether these methods and ideas are applicable to other materials. Magnesium oxide single crystals, in common with most crystals of the rock-salt structure, deform plastically but fail by cleavage after a small strain when tested in tension. It was hoped that larger strains would be attained using tension-compression. There is, in addition, evidence 8a which shows that slip on the probable cross system, (001) < 110>, is difficult in magnesium oxide; it may therefore be possible to attain high stresses by strain hardening. 1) EXPERIMENTAL PROCEDURE Experimental methods used in this study were based in part on techniques reported in papers of Stokes, Johnston, and Li.&apos; MgO blocks, purchased from Norton Co., were used without further annealing. Specimens were cleaved to dimensions approximately 0.125 in. sq and 1 in. in length. The gage section, formed by chemical polishing, was sprinkled with 280 mesh silicon carbide particles in order to introduce fresh dislocations. The crystals were then cemented into cylindrical aluminum adapters and clamped in an Instron testing machine. One of two alternating straining programs was used. In the first, total cross-head travel was established and increased in steps after various numbers of cycles. In the second, a capacitance gage was used to directly measure the elongation of the specimen and the crosshead was controlled so as to keep the plastic strain amplitude constant. The straining was always symmetrical with respect to the initial, zero strain condition. While both procedures produce strain hardening, only the latter permits a measure of the total plastic strain so that hardening curves may be drawn. Constant plastic strain amplitude tests were done

Jan 1, 1963

By Franklin Moeller

Considering the really short time that has elapsed since hydro-metallurgical processes of extracting copper from ores have been extensively developed, and the large scale on which this method is practised at Ajo, the successful operation of the plant must excite the admiration of every visitor. The application of the mechanical unloader to the purpose for which it is employed at Ajo is but an extension of the use for which it was originally designed, some 20 years ago, for unloading iron-ore vessels at the lower lake ports. At a time when the maximum output of a single machine, with men shoveling the ore into buckets, did not exceed 40 tons per hour, G. H. Hulett invented and constructed a mechanical unloader with a bucket of 10-ton capacity. This machine was of the steam hydrauIic type, and it is still in operation after nearly 20 years of service. Several more machines of the same type were built, but the difficulties of power transmission, and the rapid development of electrical machinery, soon led to the adoption of electrically operated machines, and practically all of the later installations have been thus equipped. This electrically operated unloader has proved so well fitted for the work that the design of the lake boats has been practically revolutionized, with special view to the ease and rapidity of unloading. It is now possible to unload a 12,000-ton vessel with four machines in a little over 2 hr. This has necessitated an increase in the capacity of the machine, and the largest yet built handles an average load of 17 tons, with a possible maximum of 20 to 21 tons. The Ajo excavator is somewhat simpler in design than the iron-ore unloader, in that it is not provided with a larry. In the unloader, the iron ore, picked up by the bucket, is delivered to a larry or transfer car which runs in the lower chord of the bridge; the larry then delivers the ore either into railroad cars, standing on several parallel tracks, or into the stock pile, thereby saving the time that would be required for the bucket to make the trip. At Ajo, the bucket delivers the tailing

Jan 1, 1920

By J. Larsen-Badse

The resistance to dry sliding abrasion us been measured for some hardened and tempered carbon steels. The wear resistance, defined as the reciprocal of the volume loss per unit sliding distance, is found to vary linearly with the hardness of the abraded surface. The data fit two lines, one, of high slope, for the pure metals and for the annealed steels and another, of smaller slope, for the hardened and tempered carbon steels. A relationship is developed showing the alwasion resistance to be proportional to the hardness of the material and to en, where n is the exponent of work hardening. The values of n and of the hardness FOR most materials a high resistance to abrasion is generally associated with a high value of indentation hardness. Brine11 even used the abrasion resistance as a measure of hardness in an attempt at expanding the hardness range covered by his indentation-hardness test.&apos; When comparing different types of materials there is, however, only a very general relation between abrasion resistance and hardness. Sample brittleness seems to contribute to the deviation from a direct relationship. In general, increased brittle- are those relating to the heavily deformed abraded surface. However, for materials having n constant or varying m a regular manner over a wide range of strailts, the wear resistance will also be a linear function of the bulk hardness of the material. For tempered steels containing a dispersion of cementite in ferrite the increase in wear resistance due to the dispersion is found to be a linear function of the volume fraction of Fe3C and of the reciprocal of the mean free ferrite path. This indicates indirectly that the yielding of this material, when it is already heavily deformed, follows the Orowan criterion. ness leads to decreased abrasion resistance.&apos; When comparing materials in the same group, e.g., the metals, the spread is less. Kruschov found that the abrasion resistance of pure metals and of annealed steels is directly proportional to the hardness of the metal. For hardened and tempered steels he obtained another linear relationship, one of smaller slope and intersecting the line for the pure metals at the location of the data for the annealed steels. He also found that the abrasion resistance of cold-worked metals is independent of the degree of cold working. This shows, as he concluded, that the process of abrasion work hardens the metal in the surface region to a very high degree.

Jan 1, 1967

By H. D. Strain

Potash is a generic term used to describe a number of compounds containing the element potassium (K), which is one of the three major plant nutrients. Potash content of these compounds is commonly expressed as percentage potassium oxide (K20), and the market value of the products is usually dependent on contained K20. Most of the potash produced (approximately 95%) is used in agriculture as fertilizer. Potash products today are shipped by producing countries all over the world to meet the need for increased food production. WORLD POTASH RESERVES Apparent world potash reserves are shown in Table 15.2B.1 and total some 96 billion mt of recoverable potassium oxide. The Canadian reserves account for approximately one-half the known world reserves. The reserves shown represent the potassium oxide content of a variety of potash-bearing ores. The major ore is sylvinite (a mechanical mixture of KC1 and NaCl) from which KC1 or sylvite is recovered (63.2% K20 equivalent). Other ores, such as langbeinite (potassium, magnesium sulfate-22.7% K20) and carnallite (potassuim magnesuim chloride-17.0% K20) are also mined in various parts of the world, mainly for the potassium content. Theoretically, potassium is available in almost unlimited qualities from brines and saline waters. To date, the cost of recovery from salt brines has been greater than the costs by mining operations, but new technology could make brines and seawater an increasingly important source. POTASH PRODUCTION Potash production dates back to the 1860s in Germany, which was the only recorded producing country until 1912. By 1912 Germany was producing approximately 1 million tpy of K2O equivalent. Production began in France in 1913 and in the United States in 1915. World potash production and consumption are shown in Table 15.2B.2 for the years 1950-51 through 1967-68. In 1967 and 1968 the major potash-pro¬ducing countries in order of importance were those listed in Table 15.2B.3. By 1972 Canadian production had outstripped U.S. production. WORLD CONSUMPTION World consumption of potash for the years 1950-51 through 1967-68 is shown in Table 15.2B.2. Table 15.2B.4 shows total world consumption of potassium materials in 1967-68 by country. More than 70% of the world&apos;s potash is

Jan 1, 1976

By D. N. Collins, J. R. Stevens

New expressions in the technical assessment of mineral concentration processes have been defined as Concentration Efficiency: Concentration Index: where h is the percentage of head grade; c, percentage of concentrate grade; maximum percentage of concentrate grade obtainable at 100 pct recovery; and R, percentage of recovery. In developing these expressions, the basic requirements for the technical assessment of concentration operations have been defined and the various expressions suggested by others for asses-sing the results of such operations reviewed and criticized. The derived expression for E, depends on a simple combination of recovery (R) and a new term selectivity (s), where S is (c-h)/(c,,, -h), by definition. Limitations of the new expressions are discussed and examples given of applications to concentration operations. It is also shown that the formulae for E, and I, can be applied to scatter diagrams, multi-component systems, size-assay analyses, and classification data. Many methods of quantifying concentrating operations have been proposed. These include direct presentation of weights and assays together with the derived data for other items such as ratio of concentration, recovery, rejection losses, and enrichment ratio. These terms are now accepted mineral dressing nomenclature, and in the assessment of the results of concentration processes, concentrate grade and recovery are together the most widely accepted measures of concentration efficiency. For practical purposes when the concentrate grade is equal to or greater than a predetermined acceptable value, test or operation results may be assessed by the recovery figures alone. Similarly when recovery is within specification, the operation can be assessed by concentrate grade. The ultimate assessment and selection of a concentration operation is—admittedly with a few ex-ceptions—a matter of economics. There is however an intermediate stage in laboratory and pilot plant testing when evaluations must be made without the assistance of an economic yardstick. Numerous techniques are employed to facilitate such evaluations among which may be mentioned graphical plots of grade vs recovery, cumulative weight vs cumulative recovery, scatter diagrams, etc. The desirability of combining metallurgical results and product weights and assays into a single value has been recognized and discussed by many workers1-" with the resultant suggestion and use of such terms as Metallurgical Efficiency ,&apos; Summation Index,&apos; Selectivity Index,&apos;- and Absolute Efficiency B None of these expressions completely satisfies all the requirements of a selectivity or efficiency formula and it is proposed in this report to define the terms, relating to the technical efficiency of a concentrating operation, concentration index and concentration efficiency. Examples are given of their application and a comparison is made of their use in relation to other formulae. All product grades in this paper will be referred to in terms of mineral percentages and a list of symbols is given in Table I. CONSIDERATION OF EXISTING FORMULAE A number of the existing efficiency and index terms used in quantifying concentrating operations are considered below. Metallurgical Efficiency:

Jan 1, 1961

The Ashanti mine is located in the Gold Coast region of West Africa in Ghana, where gold is found over a wide area of about 105,000 sq km (40,000 sq miles). Ashanti has been the most productive of all the lode mines and has been one of the world&apos;s notable gold producers for over 50 years. The mine is located at Obuasi, 113 km (70 miles) north and east of Tarkwa, which is 80 km (50 miles) inland from the coast. The Ashanti lode is worked on 300 m (1,000 ft) of strike length over a width of 12 m (40 ft), and by 1934 it had attained a depth of 700 m (2, 300 ft). The lode is in pre-Cambrian granite intruding a system- of shists, phyllites, etc., containing fragments of the wall rock with quartz, carbonates, sericite and small amounts of pyrite, arsenopyrite, sphalerite, galena, and gold. The shattered quartz fractures carry sulfides and gold, most of which are invisible. (William Harvey Emmons, "Gold Deposits of the World," McGraw -Hill, 1937.) Because the mill treats an ore composed of refractory sulfides - chiefly arsenopyrite and fine free gold - the flowsheet reflects the complex ore characteristics. Plane tables are used in the grinding, circuit for free gold recovery. After roasting the sulfide flotation concentrates, additional free gold is recovered from the calcines with a rotary belt concentrator and the two concentrates are combined for feed to the amalgamation pans, with the resulting gold amalgam being retorted. Following four-stage crushing to minus 9. 5 mm (318 in.) with two stages of grinding, the cyclone overflow is floated for sulfide recovery. The flotation concentrate is roasted with the calcine reground and roughly cyanided to scalp out the gold readily dissolved in one of two cyanidation circuits. The cyanide residue joins the thickened flotation tailing for cyanidation in a complete CCD circuit using filters. The gold precipitates are calcined before melting in pot-type furnaces. The tonnage to the mill of 3,175 mtpd (3,500 s tpd) with a content of over 14.7 g per mt (0.43 oz per st) places the Ashanti production as one of the top producers in the world; and with its past long life, the total production rates it as one of the world&apos;s great mines.

Jan 1, 1981

By R. L. Stone

None of the currently used methods of testing bentonites as binders for strengthening iron ore pellets is entirely adequate by itself. Since more knowledge was needed about 1) which properties of bentonite most affect pellet strength, and 2) the precise mechanism by which strengthening is produced, attention was paid to possible new techniques for rating bentonites. With the structural nature of bentonites — in particular, the montmorillonite component — the possibility of rating bentonites on the basis of zeta potential was investigated. These bentonites were also rated by conventional tests and measurements, including colloid content, barrel yield, pH, chemical analysis and physical tests on green and dry pellets. A mechanism for pellet bonding based on electro-kinetic forces between the bentonite and the iron ore concentrate appears to be feasible. In pelletizing, from 6 to 18 lbs of bentonite is added per ton of concentrate to give strength to the green and dry pellets. In some operations even more bentonite than is required for strengthening purposes must be added for moisture adsorption to pennit balling of the concentrate. Since bentonite is thus a major item in pelletizing, more knowledge about the way in which bentonites act as binders and better ways of rating this ability would help in the difficult job of selecting the right bentonite for a given concentrate. The fact of the matter is that although several methods are in current use for testing bentonites, none of these methods are consistently foolproof. For example, a given bentonite may rate high by several standard tests yet turn out to be a poor binder, and of course the reverse also occurs. Actually, more knowledge is needed about 1) which properties of bentonite most affect pellet strength, and 2) the precise mechanism by which strengthening is produced. In this study attention is therefore paid to possible new techniques for rating bentonites. With the structural nature of bentonites as a starting point, we investigated the possibility of rating bentonites on the basis of the zeta potential, i.e., the electrokinetic ootential on the montmorillonite colloids — the main component of bentonite binders. The zeta potential technique was applied to a range of commercial bentonites from several areas in the U.S. and abroad. These bentonites were also rated by conventional tests and measurements, including:

Jan 1, 1968

By J. Szekely, Robert G. Lee

Calculations are presented for the prediction of the combined radiative-convective heat loss from molten steel held in a ladle, covered by initially molten slag. A mathematical formulation is given and numerical solutions are presented of the resultant partial differential equation. It is shown that a slag layer of about 2 to 3 in. thick would prevent any significant heat loss from the top metal surface for Periods of up to 1 hr. For thinner slag layers (or in the absence of the protective slag) appreciable heat losses would occur; plots given in the paper allow the quantitative evaluation of the fall in the metal temperature due to this heat loss. The ironmaking-steel processing operation involves several intermediate steps where the molten metal is held or transfered in ladles. The quantitative assessment of the heat losses occurring under these conditions is of considerable importance in process design considerations since the temperature of the metal may have to meet rigid specifications in any given part of the processing sequence. While the metal is held in a ladle, heat is lost by two mechanisms: i) by conduction into the ladle walls; ii) by radiation and convection from the top surface. Calculations relating to the conductive heat loss, are readily made and information may be found in the literature both on data pertaining to metallurgical situations and on the techniques that are available for performing additional computations.&apos;-3 The evaluation of the combined convective-radiative heat loss is less straightforward, especially when there exists a protective slag layer covering the metal. In this latter case, as the top slag surface loses heat partial solidification may occur and the latent heat thus given up by the slag may represent an effective barrier to the heat loss from the metal. A semiquantitative assessment of the role played by the slag in preventing heat loss from the metal has been made in a previous paper,4 where it was shown that a slag layer 6 in. thick would act as a near-perfect insulator for periods of up to 2 hr. However, this first paper reported on an essentially preliminary investigation that considered only one slag layer thickness; furthermore, the boundary conditions used for the calculations were somewhat restrictive since no allowance was made for changes in the metal temperature. The purpose of this second paper is to extend the scope of the preliminary investigation previously reported by: i) considering several slag layer thicknesses; ii) allowing for variations in the metal temperature; and finally iii) by performing calculations on the net loss from the metal rather than from the slag surface. FORMULATION Consider a slag phase extending from y = 0 to y = L1, covering a metal phase that occupies the region extending from y = Ll to y = L, as illustrated in Fig. 1. Denote the slag and metal temperatures by TS and Tm and assume that initially both slag and metal are molten, having a uniform temperature Ti. At time = zero the surface represented by the y = 0 plane (top slag surface) is brought into contact with cold air, the temperature of which is given as T Thus for time > zero, heat will be transfered from the slag to the air by natural convection and thermal radiation; as a result of this heat loss the slag temperature will fall and after some time a solid phase is formed and a solidification boundary (more realistically described as a zone) will move progressively from y = 0 toward y = L1. Once the region of temperature gradients, moving ahead of the solidification zone, reach the slag-metal interface (y = L1) there will be a net transfer of heat from the metal through the slag to the cold air, across the y = 0 plane. In the formulation of moving boundary problems, involving phase changes that require computer solution, it is convenient to represent both phases by a single equation, making allowance for the latent heat of solidification by assigning an appropriate temperature dependence to the specific heat content. Thus the energy equation for the whole of the slag may be written as follows:

Jan 1, 1969

By R. W. Bartlett

Hydrometallurgical processes for copper flotation concentrates avoid the SO2 emission problems associated with smelting, but they require oxidation of copper to obtain solubilization during leaching. The resulting copper ions must be reduced to metal, which normally requires large amounts of energy or energy equivalent reagents such as hydrogen or iron. The conventional reduction method is electrowinning which is labor intensive and involves 35 to 40% of both the entire operating and capital cost of the better hydrometallurgical processes. There is considerable incentive to reduce the cost and energy required for producing copper ions from solution. Furthermore, it is thermodynamically possible to do this using sulfur contained in the concentrate itself. In fact, the thermodynamic potential of oxidizing sulfur from S+4 to the S+6 state is greater than required to reduce cupric ion to metallic copper.1 Realization of this potential in a practical process would significantly reduce the cost of hydrometallurgical processing of sulfide concentrates because sulfur would become a cost-free reducing agent. An apparently practical method is to use SO: and HSO3 ions, obtained by ammonia scrubbing of SO2 from a concentrate roaster gas, to reduce leached copper to metal in an autoclave. This reduction step is the Jumau 1907 A series of steps, including sulfite reduction, constituting a crude flowsheet was suggested by Parker in 1976.3 The present paper describes a detailed complete. sulfite process flowsheet based on materials and energy balances, discusses the chemistry, and discusses further work that is required to develop this process. Process materials, utilities costs, and energy requirements are projected to be unusually low because of the utilization of sulfur from the concentrate. Process Description The sulfite reduction process requires a copper concentrate containing approximately as much sulfur on a weight basis as copper. For those few concentrates that are deficient in sulfur, additional sulfur can generally be obtained by allowing more pyrite to float in the concentration process. The process steps, shown in [Fig. 1], are (1) a fluid bed roast of the concentrate, (2) a leach of the roaster calcine using recycled barren liquor from the reduction autoclave, (3) scrubbing of the SO2 gas with ammonium sulfite liquor to produce ammonium bisulfite, (4) reduction and precipitation of copper in an autoclave by combining part of the SO2 scrubber effluent (SO5

Jan 1, 1980

W: 0. HOTCHKISS, Madison, Wis. (written discussion*).-Mr. George&apos;s paper on. "The Wisconsin Zinc District" is a very complete and clear description, which it is a pleasure to read. His discussion of the origin of the ores is the least satisfactory part, and Mr. George correctly states present knowledge when he begins his paragraph on a origin with the words "the general supposition is, etc." There is still much to be learned regarding the details of the processes by which the ores have reached their present situation. In this the Wisconsin district differs not at all from other mining districts. It is my hope that Mr. George and others, equally clear-thinking, who are in close touch with these operations will be able at an early date to give us more information on this subject.

Jan 3, 1918

By R. G. Stickney

INTRODUCTION The Basalt Waste Isolation Project, conducted by Rockwell Hanford Operations under contract to the U.S. Department of Energy, is assessing the feasibility of the disposal of commercial radioactive spent fuel/high-level waste in deep geologic formations in basalts beneath and within the Hanford Site in southeastern Washington State. The Basalt Waste Isolation Project mission is to identify and characterize a potential geologic repository site in basalt and to develop associated facilities and technology required for the potential permanent isolation of radioactive wastes in basalt formations. In support of the Basalt Waste Isolation Project, a laboratory testing program was conducted to evaluate the performance of the Council for Scientific and Industrial Research&apos;s "Doorstopper" biaxial strain cell. To support site characterization studies and the design of repository openings, in situ stress measurements will be conducted within the reference repository horizon (~970 m). The Doorstopper is one of the instruments being considered for use in these measurements. It was selected because it requires only a small volume of work to achieve stress relief, which is important in a closely jointed basalt. Also, test durations are short, permitting a large number of measurements to be made within a given time to increase confidence in the results. Within the reference repository horizon, temperatures are expected in the range of 40º to 60°C with a humidity of up to 100%. For this testing program, test conditions were selected to closely resemble those expected within the repository horizon. The Doorstopper test, mainly developed by Leeman (1964), is a strain rosette relief method of measuring rock stress in situ. The Doorstopper consists of a four element, 45º strain gage rosette

Jan 1, 1984