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By J. B. Ackerman, J. D. Miller

Alunite [KA13(S04)2(OH)6] is a promising non-bauxitic aluminum resource, the domestic reserves of which are estimated to be 800 x 106 tons at 35 percent alunite. The major gangue mineral associated with alunite is quartz. The thermochemical process that has been developed for treatment of this ore does not provide for the initial removal of gangue prior to processing. In this regard, it may be desirable to concentrate the ore to gain significant technical and economic advantages. Experimental results indicate the effectiveness of an oleic acid flotation technique and exemplify the problems encountered in fine particles processing. These results show that the quality of separation studied as a function of temperature, collector addition, and depressant addition is dependent on the degree of liberation achieved during size reduction. At coarse grinds the quality of separation was limited by the presence of locked particles while for finer grinds where more extensive liberation is achieved, better separations were obtained. The degree of liberation for each size interval was determined by an SEM/energy dispersive x-ray technique and found to be independent of the extent of size reduction, suggesting that liberation occurs by indiscriminate fracture and not by detachment. Up to 89 percent liberation was realized for a minus 400 mesh grind. Under these circumstances, the quality of separation appears not only to be limited by the remaining locked particles, but, also, by contamination of the alunite concentrate with fine quartz gangue (< 2µm). The best separation for the minus 400 mesh grind was a rougher recovery of 88 percent with a grade of 65% alunite at 50°C.

Jan 1, 1980

By Per-Martin Sandgren, Alrik Anttila

LKAB&apos;s ores have specific mineralogical properties that make them especially suitable for the production of supergrade concentrates. Conditions are particularly good for this purpose at Malmberget, where a concentrate for high-purity sponge iron production has been produced for many years. This product assays approximately 71.5% Fe and 0.20% SiO2. Potential exists for still further improvement. In recent years, the production of fractioned concentrates has been started. The material lies within the size range 74-125 µm and the approximate chemical analysis is 71.4% Fe and 0.30% SiO2. Other special products are under discussion. Malmberget ores are also suitable for the production of high- purity direct reduction pellets with a SiO2 content below 1%.

Jan 1, 1983

By D. M. Wade, F. F. Redfield

PRODUCTION at Aguilar started in 1936 and by 1942 the mine had reached a peak of 25,000 tons a month. Present production is only about 60 per cent of this capacity because of difficulties in railroad transportation and the rationing of fuel oil brought about by the war. In the early years, cut-and-fill stopes were tried, but square sets were found to be the most practical method in the southern section of the mine, where blocky, treacherous ground and greater mining widths are prevalent. Square sets and opencut mining were the rule until 1942, when it became increasingly difficult to get mine timber. The ore beds in the northern section of the mine were found amenable to cut-and-fill stoping as the rock was firmer and the veins narrower. At present 69 per cent of the ore comes from cut-and-fill stopes, 27 per cent from opencut, and only 4 per cent from square-set stopes. Massive galena and sphalerite with comparatively little quartzitic gangue constitutes the higher-grade ore. The gangue minerals that can be detected by eye in low-grade ore are rhodonite, wollastonite, and pyrrhotite. Wall rocks range from shale to quartzite, with occasionally a little limestone.

Jan 1, 1947

IT is customary to make an alloy of copper in the same way, not to increase its quantity as with gold or silver, but to corrupt it for the art of casting and to destroy a certain natural viscosity in it. For this result it is accompanied by tin and also, sometimes, brass. But its proper and true alloy is fine tin, except, however, when you wish to make hammered works. For these, copper must be pure and without any trace [of impurity], otherwise it cannot be hammered thin, gilded in the fire, drawn into a wire, or made into vessels for use as is customary; but in casting, tin is almost essential. When you come to this association, just as it changes nature and appearance according to the proportion of tin which you give it, so it changes name and is no longer called copper but bronze, or, for a greater distinction of the quantity of tin for every hundred or more of copper, it is called by skilled workmen metal of more or less fineness as it contains more or less tin. This is recognized by the whiteness and brittleness: changing from red, which is the color of copper, it becomes white; from soft and flexible it becomes hard and brittle as glass. This admixture removes it so far from its original nature that one who does not know that it is a compound material believes it to be one of the number of metals engendered by Nature. Now, so that you may understand well, I tell you that various kinds of this are made, all according to the proportions of the quantities and according to the kind of work. For one kind of bronze is desired by those who make statues; this is not desired by those who make artillery or by those who make bells, mortars, basins, and similar cast objects. Now in order to alloy it into the species of bronze, eight, nine, ten, up to twelve pounds of tin are put with every hundred pounds of copper. Those who want it to make bells put twenty-three, twenty-four, twenty-five, and twenty-six, depending on the tone, on whether they are large or small in form, and on whether they wish to make the tone deep or sharp and clear. From twelve upwards is used for all other works that need it, either for hardness or to give them fluidity in casting by surpassing the degree of

Jan 1, 1942

By John M. Hartman

On May 19th, 1857, an English patent was granted to E. A. Cowper for heating air or other gases under pressure by means of a regenerator inclosed in an air-tight iron case, having between the regenerator and case a lining of brick. This patent provided for heating the stoves by a separate fireplace, or by gas direct from the blast furnace. A number of forms of interior arrangement of the brickwork are shown in the drawings; also hollow poppet valves with hollow stems, and a pipe inside of the stem for circulating the water; the valve seats have coil cast in them for water circulation to keep them cool; slide valves, with snake coil cast in the discs, are shown, and the use of cold air for cooling the valves is also described. The combustion-chamber of these stoves was central, and openings mere provided at the top and bottom to get into the stoves. These Cowper stoves are all circular in section. November loth, 1865, an English patent was granted to Thomas Whitwell for regenerative stoves for heating air or gas, provided with cleaning openings at the top and bottom capable of being closed with firebrick plugs and doors. The drawings show a rectangular stove inclosed in an iron case. The interior brickwork has numerous upand-down passages through the stove, but there is no claim on the interior construction. March 3d, 1868, an English patent was granted to Charles Cochrane for a slide valve to be subjected to high heats. The disc of this valve was hollow, and had a circulation of water through it by the two hollow stems that operated it. The valve-seat was detachable and had a coil cast in it, through which water circulated. The valve and seat were placed on an incline to the body to cause the valve-disc to lie on the valve-seat. A cap was placed on the bottom of the body to get at the interior readily. January 5th, 1870, an English patent was granted to Siemens, Cowper and Cochrane for the construction of regenerators in firebrick stoves, with numerous vertical passages of sufficient size to allow a brush to pass through and clean them. These passages had slight projections on the sides to tarn the air over and over as it passed through. A claim also covered the use of horizontal passages connected at each end alternately, and the use of blasts or jets of air,

Jan 1, 1880

By A. Tobey Yu

Despite its inherent low Btu content, western coal&apos;s low sulfur content may provide a near-ideal answer to satisfy US goals for clean air and energy self-sufficiency. Vast reserves, amenable to low-cost strip mining, serve as a powerful economic basis for its viability as a primary and dependable source of energy for the coming decades. A major consideration, the solution of which may not be obvious, is the transportation cost. These deposits are remote from the utilities of the heavily populated eastern regions of the US (Fig. 1). A coal mine in Montana is likely to be more than 1500 miles from its utility market, while an eastern coal mine may be situated only 200 miles away. Thus, the question raised is: how can the haulage cost of western coal be substantially reduced to an acceptable level?

Jan 7, 1975

By Joan M. Hartman

ON May 19th, 1857, an English patent was granted to E. A. Cowper for heating air or other gases under pressure by means of a regenerator inclosed in an air-tight iron case, having between the regenerator and case a lining of brick. This patent provided for heating the stoves by a separate fireplace, or by gas direct from the blast furnace. A number of forms of interior arrangement of the brickwork are shown in the drawings; also hollow poppet valves with hollow stems, and a pipe inside of the stem for circulating the water; the valve seats have coil cast in them for water circulation to keep them cool; slide valves, with snake coil cast in the discs, are shown, and the use of cold air for cooling the valves is also described. The combustion-chamber of these stoves was central, and openings were provided at the top and bottom to get into the stoves. These Cowper stoves are all circular in section. November 10th, 1865, an English patent was granted to Thomas Whitwell for regenerative stoves for heating air or gas, provided with cleaning openings at the top and bottom capable of being closed with firebrick plugs and doors. The drawings show a rectangular stove inclosed in an iron case. The interior brickwork has numerous up-and-down passages through the stove, but there is no claim on the interior construction. March 3d, 1868, an English patent was granted to Charles Cochrane for a slide valve to be subjected to high heats. The disc of this valve was hollow, and had a circulation of water through it by the two hollow steins that operated it. The valve-seat was detachable and had a coil cast in it, through which water circulated. The valve and seat were placed on an incline to the body to cause the valve-disc to lie on the valve-seat. A cap was placed on the bottom of the body to get at the interior readily. January 5th, 1870, an English patent was granted to Siemens, Cowper and Cochrane for the construction of regenerators in fire-brick stoves, with numerous vertical passages of sufficient size to allow a brush to pass through and clean them. These passages had slight projections on the sides to turn the air over and over as it passed through. A claim also covered the use of horizontal passages connected at each end alternately, and the use of blasts or jets of air,

Jan 1, 1880

By John E. Carr

A bewildering variety of continuous thickeners and clarifiers are used in mineral industries and other heavy industries. General topes arc: ? Conventional thickeners ? Thickeners with flocculating feedwells ? Solids contactors ? Tray thickeners ? Lamellas ? Conventional thickeners with Lamellas ? High flow, fluid bed thickeners These units must be adapted to varied applications. They must provide operating flexibility to meet process needs. Such factors multiply the decisions which must be made prior to the selection of a thickener. The seven types of thickeners listed above, with their variations, share design concepts which limit their application and which can be used to evaluate their performance. Some of these arc: ? Feed velocities must be limited and/or controlled. ? Natural or reagent flocculation is necessary. ? Horizontal settling area is the basis for capacity. ? Overflow velocities must be limited and. or controlled. ? A controlled underflow is necessary. Thickener selection is the application of these needs to the basic process concepts. The basic function of a thickener, in the minerals industry, is the separation of a thickener sludge containing fine solids from water. In addition to this primary separation function, others include: ? Separation of solids from water ? Separation of flocculated solids from slimes ? Storage of solids ? Control of percent solids ? Storage of water ? Process control by absorbing fluctuations in feed ? Process control by increase in head clue to the depth and percent solids of the thickened sludge. Most thickener designs, suitable to minerals applications, are based on the separation of minus 65-mesh, 2.7 specific gravity solids containing slimes. These designs are routinely adapted to coarse and or denser materials. Materials of lower specific gravity such as coal, wood, char, some chemical precipitates and organics act Hike slimes if a thickener is used. Perhaps the most conservative approach to sizing and application is the use of comparative data from similar plants or similar industries. Thickeners are-routinely used in a great number of industries from which there is good historical information available. In the minerals industry, thickeners are used in: clay sizing, coal refuse, heavy media, copper concentrates, copper tailings, gold flotation concentrates, and countercurrent decantation for cyanidization in gold and CCD in uranium. Concentrate and tailings thickeners are used in iron ore plants, and concentrate and tailings units in lead-zinc plants. In addition, there arc thickeners designed to handle lime mud; limestone concentrates and tailings; phosphoric acid; potash; and rutile concentrates. Thickeners are, in addition, widely used in gravel plants for water reclamation. This gives excellent data on clay settling in some cases. Thickeners are used extensively for slurries generated by flue dust from wet scrubbers. Thickeners in flue dust applications are operating with considerable variations, treating blast furnace dust. BOF flue dust, and wood char. Similar applications are met in the power industry with fly ash and SO2 scrubber systems. Even though field data is preferred, such data has limitations as listed: ? Flowsheets are seldom identical. ? Data is often incomplete or limited. ? Ores vary significantly. ? Operating needs are changing. For these reasons, mathematical and theoretical methods are needed for application of field data to a specific project. In addition, laboratory data for similar projects and, if possible, laboratory data on a representative sample should be used to specify the thickener. There is a great volume of thickener literature. The references for this paper, along with their references, and past AIME papers provide excellent background. Older literature can be reviewed in light of current methods. It is interesting to note that the requirements for higher flow rate designs, more extreme flocculation systems, etc., can be logically introduced by reading Gaudin&apos;s Sections on Flocculation and Thickening, which was written before field installations could illustrate all of his ideas. The problem, then, is to select hardware which will take into consideration the functions of flocculation, separation of the solids, water clarity, control of the underflow, and control of the overflow. Reliable operation and simple control are customary.

Jan 1, 1979

The income of the Institute is derived mainly from dues, advertising in MINING AND METALLURGY and sale of publications. These sources are fortunately supplemented by the interest from invested funds now amounting to over $500,000, a considerable portion con- stituting endowments for especial purpose?s. Aside from the medal funds the principal endowments are the James Douglas Fund for support of the Library, the Rocky Mountain Fund and the Seeley W. Mudd Memorial Fund. The income from the two last named is available for the support of research and a variety of other purposes. It is allocated by the Board of Directors upon recommendation of standing committees in each case. An active effort is being made to increase the endowments of the Institute, it being felt that unusual opportunities exist for research and public and professional service through the organization.

Jan 1, 1937

By R. W. Jenkins

TRACKLESS equipment is being used by the Coplay Cement Manufacturing Co. to develop a folded limestone deposit economically and safely with inexperienced men. Cost and quality of development stone results in a saving when it is used in- stead of outside sources of limestone. This favorable cost has been attained by following a methods development and training program described in this paper. It is expected that further savings in favor of the mined stone will be made when stopes now being prepared are in production. The present equipment will be used for loading from stope draw points. The limestone is blended with quarried cement rock to serve as raw material for Portland cement, first produced in this country by Coplay Cement Manufacturing Co. from a process developed by its president, David 0. Saylor, in 1873. For many years the cement rock alone or with a small amount of added limestone was used for raw materials. Later as specifications were raised and the richer portions of quarries were depleted, limestone from outside sources was required in increasing amounts. In 1925 D. J. Uhle, the general manager, began to prospect company property for limestone.

Jan 1, 1954

By John V. Beall

AMONG electrical manufacturers, interest in the outlook for copper supplies has not diminished despite signs of a healthier market following price decontrol. Reason: Electrical utility industry output is expected to grow at an average annual rate of 6 pct for the next 15 years. A more graphic description of the growth ahead is given in the Eighth Quarterly Report of the Director of Defense Mobilization as follows: "Power generating capability, which amounted to 63 million kw at the time of Korea, now stands at more than 81 million. The established goal for installed capability at the end of 1955 is 116 million kw. Orders already on the books of power equipment manufacturers, if fulfilled on schedule, will bring installed capability to over 123 million kw by the end of 1956-and more orders for equipment are yet to come." These statistics portend an impressive requirement for copper, not only for power plants and distribution lines themselves but for the multitude of industry and domestic end-uses.

Jan 4, 1953

By A. G. Guy

Thermodynamic activity rather than chemical composition is basic to the analysis of diffusion. This is the essential conclusion reached by Darken1-3 and by Birchenall and Mehl.4 If so, it is reasonable to expect that diffusion calculations should be carried out using activities. This paper will illustrate a method of employing activities in analyzing interstitial diffusion. For interstitial diffusion (of carbon in iron and probably generally) the diffusion of the solvent atoms, volume changes, and other such complexities not explicitly provided for in the usual diffusion equations cause a negligible error of the order of 1 pct. Therefore it is relatively easy to test the validity of an analysis based on activities in the instance of interstitial diffusion. The mathematical treatment of the more complex substitutional diffusion could be simplified once the adequacy of the activity method had been established for substitutional diffusion. Thus the present work has the purpose not only of providing a superior analysis of interstitial diffusion but also of indicating a possible approach to the more important question of an adequate treatment of substitutional diffusion. Development of the Diffusion Equation Although activities are convenient quantities to determine experimentally, they are derived from the more fundamental fugacities.9 Possible confusion in using activities can be avoided by basing equations involving their use on the relationship, dP = -D1d/ax dt [I] where, dP = the number of grams of solute crossing one cm2 in the time dt sec. D, = the diffusion constant for use with fugacities; it is assumed to be constant at a given temperature, and has units of secs. f = the fugacity of the solute in the solid solution; the units are those of pressure, dynes per cm2. x = the distance in cm. Eq 1 can be considered to be the basic form of the first Fick Law for one-dimensional diffusion. In order to convert Eq 1 into an equation involving activities, a choice of a standard or reference fugacity, f°, must be made: then. a = f [2] where a is the activity corresponding to the fugacity f. However, the value of a is also determined by the relation, a = a\c [3] where a is the activity coefficient and c is the concentration. For a given value of a (such as a = 1) it is evident that the units used for expressing concentration will affect the numerical value of a corresponding to a given amount of carbon dissolved in y-iron. Since weight per cent concentration is so widely used, this unit will be adopted as the standard in this paper. When the value off given by Eq 2 is substituted in Eq 1, the first Fick Law becomes, dP= -D,f°dl [4] Since D,f° is constant (at a given temperature), the second form of Fick&apos;s Law can be obtained in the usual manner, -dt D&apos;f° dx [5] where c&apos; is the concentration in g per cm3. In terms of weight per cent concentration, c, Eq 5 becomes, dc _ 100 d2a Tt = T /f w [6] where p is the density of the solid solution and will be assumed to be constant. In order to simplify Eq6, it is necessary to consider the standard state, f°. Standard states are chosen for convenience. There appears to be one especially convenient choice for the analysis of interstitial diffusion. If the activity coefficient, a, is set equal to unity at infinite dilution of the solute, 100 then — 100D1fl° is approximately equal P to the usual diffusion constant in dilute solutions. This is true since the activity, a, in Eq 6 can be replaced by the concentration, c, with little error in an infinitely dilute solution.* Here f1° is the standard fugacity necessary to achieve this standard state. It is proposed that a diffusion constant Da&apos; be defined, Da1=100/pD1f1° [7] where A is the metal whose diffusion is * The second derivative of a is equal to the second derivative of c in dilute solution only if -da/dr- is. zero. Although theory does not require this condition, in the systems C in Fe, and Zn in Cu, dc is in fact found to be essentially zero.

Jan 1, 1950

By D. E. Spiller

Introduction Precious metals recovery from ore can generally be accomplished using gravity concentration, flotation, and/or hydrometallurgical (leaching) techniques. The objective of this paper is to show why gravity concentration can be an important part of recovery systems that employ leaching as the primary unit operation. A brief discussion of modem gravity concentration equipment is also presented. Discussion Gravity concentration of ores has generated increasing interest in recent years. Reasons for this interest include: • Gravity concentration is environmentally attractive. There is little or no use of reagents. Hence, it is relatively nonpolluting. • The cost of cyanide has continued to increase. Therefore, cost savings may be realized whenever leaching feed tonnage can be reduced by preconcentration. • Compared to flotation and leaching, gravity equipment costs are low per processed ton. Field installation costs for gravity circuits usually are less because many&apos; units are supplied as self-contained modules. Also, the cost required to supply services, particularly power, to a gravity plant site are also less. In situations where preconcentration at coarse particle size is applicable, significant grinding equipment savings may be possible. • Gravity circuit operating costs are also relatively low compared to typical flotation and leaching circuits. Reagents, power, maintenance, and manpower savings in a well-engineered gravity plant may be realized. Again, if grinding is reduced, significant power and steel (media and liners) savings are possible. •In recent years, more efficient gravity concentrating devices have been developed. Benefits to Precious Metal Leaching Gravity beneficiation can complement precious metal leaching in two ways. First, the recovery of coarse liberated values before leaching may reduce leach time requirements and may reduce reagent consumption. Second, gravity preconcentration can reduce the size of a leach plant by decreasing the quantity of material to be leached. Coarse gold and silver have been shown to leach rather slowly. Kameda (1949) and Habashi (1967) have investigated the kinetics of cyanide leaching systems. They concur that in a heterogeneous reaction, the rate of gold and silver dissolution is directly proportional to the surface area. Thus, the instantaneous rate of dissolution for spherical 0.37 mm (400 mesh) gold is theoretically -25 times faster than for the same amount of gold at .841 mm (20 mesh), based on data from Fuerstenau, Chander, and Abouzeid (1979). Conversely, coarse liberated, +.841 mm (20 mesh), gold is more readily recovered by gravity concentration than is fine, -.037 mm (400 mesh) gold. Therefore, it is apparent that the two recovery systems complement one another. Figure 1 data demonstrates the potential synergism. A sample of - 3.327 mm (6 mesh) Nevada gold-bearing ore was cyanide leached using conventional bottle-roll test procedures. Gold extraction was determined as a function of leaching time. A second sample split from the same leach feed material was hand jigged to remove a coarse heavy mineral fraction, including virtually all of the +.210 mm (65 mesh) liberated free gold. This second sample, with the coarse gold and heavy minerals removed, was subjected to an identical cyanide leach procedure. Figure 1 presents the resulting comparative extraction data. Note that the percent gold extraction for the sample containing no +.210 mm (65 mesh) free gold includes the coarse gold recovered by gravity. The data show that the sample containing coarse gold required about 72 hours of leaching time to achieve 80% extraction. This compared to about 22 hours of leaching time for 80% gold recovery from the sample that contained only -.210 mm (65 mesh) free gold. Thus, there was a 69% reduction in leaching time. The improved extraction data is not wholly attributable to coarse gold removal, but rather it was the combination of gold removal and rejection of other heavy mineral cyanide consumers or leach retardants. Further investigation was not warranted at this time. Preconcentration is the second manner in which leaching systems can benefit from gravity concentration. The premise is that preconcentration can reduce the quantity of leach feed, which, in turn, may reduce leaching costs. Figure 2 presents preliminary data developed by CSMRI for US Minerals Exploration (USMX). Centennial Exploration Inc., in agreement with USMX, is proceeding with evaluations to determine the suitability of various processing schemes for recovery of gold values from the Montana Tunnels property. The data shows how the ore can be preconcentrated by gravity techniques to result in a reduced feed tonnage to secondary extraction techniques, presumably flotation or cyanide leaching. Testing has shown that Reichert cones, followed by treating the cone concentrate on spirals, can deliver about 88% gold recovery in about 13% weight, that is, 87% weight rejection. Consequently, fine grinding and reagent costs are attributable to only 13% of the plant feed rate. Cost data is not yet available, but the potential exists for significant cost savings.

Jan 1, 1985

Our examination was made in accordance with generally accepted auditing standards, and accordingly included such tests of the accounting records and such other auditing procedures as we considered necessary in the circumstances , In our opinion, the accompanying statements present fairly the financial position of the Institute as at December 31, 1951 and the results of its operations for the year ended on that date, in accordance with the method of accounting maintained by the Institute for 1951, which (except for a charge to expense of $12,500 00 in 1951, in which we concur, to provide for increasing the reserve for depreciation to an amount considered appropriate as at December 31, 1951) was consistent with that followed for the preceding year

Jan 1, 1952

By E. H. Kelton

THE purposes of this paper are to describe the use of adjustable cut and draw tools as a control test of drawing properties and to point out that no other well-known test or combination of tests determines this quality in zinc and zinc alloys. Rolled zinc has been used for making drawn articles, particularly battery cans, for many years. The development of rolling treatments satisfactory for, these purposes followed the usual course of trial and error experimentation depending upon cooperation with customers to determine what type of material was suitable. A quick test of drawing quality was highly desirable, of course. Scleroscope hardness and temper,1 having no apparent relation to drawing quality, were used simply to check the duplication of rolling treatment and grade of metal. The dynamic ductility1 test, on the other hand, was believed to be so like the drawing operation as to provide a reliable index and came to be taken as the criterion of good drawing properties. Faith in this test was strengthened by the fact that metal having the highest ductility was more satisfactory for operations, such as the manufacture of battery cups, that required several redraws. No distinction was made at that time in specifying metal as to whether it was to be subjected to simple cut and draw, redrawing or forming operations. It developed, however, that in some grades of metal the highest ductility did not coincide with the most satisfactory quality when the material was used for a single draw.

Jan 1, 1934

By David R. Mitchell

The mineral pyrite (or marcasite) occurs in coal beds as balls, lenses, veinlets and bands. Several million tons are wasted annually on the refuse dumps from coal mining and coal-preparation activities. Sporadic attempts have been made in the United States to recover this pyrite for manufacture of sulphuric acid but only a few plants have succeeded in recovering it on a commercial scale. Several chemical companies in the East became concerned about supplies of pyrite early in 1943, but by late summer of that year the situation had eased. If for no other reason than national security and self-sufficiency during wartime, methods of recovering pyrite from coal-mine refuse should be worked out and made available to coal-mine operators who are potential producers of pyrite. Eastern sulphuric-acid plants are particularly likely to have their sources of raw material blocked by war conditions, so that special studies to determine where supplies of pyrite can be obtained for these plants from inland sources are desirable. Davis&apos;s1 summarization of conditions during World War I gives the following facts: During those years, the shortage of pyrite became acute. Shipments of hand-picked pyrite were made from a number of coal mines in Illinois, Indiana, western Kentucky, Ohio and Pennsylvania. A mechanical concentrating plant for sepa- rating pyrite from picking-table refuse was built at Danville, Ill., and operated for several years, and a small pyrite-concen-trating plant treating washery refuse near Gillespie, Ill., operated for several months in 1918. The Consolidation Coal Co. of Fairmont, W. Va., constructed and operated a sulphuric-acid plant using pyrite recovered from coal-mine refuse. The U. S. Bureau of Mines during 1917-1919 made a comprehensive survey of pyrite-production possibilities from coal mines and estimated that there was a potential yearly production of 435.000 tons from Ohio and Pennsylvania and 928,000 tons from Kansas, Missouri, Iowa, Illinois, and Indiana mines. This estimate included only mines that could produce one per cent or more of their coal production as pyrite containing a minimum of 40 per cent sulphur. PyRite Production According to the Minerals Yearbook,2 approximately 660,000 long tons of pyrite was produced in the United States in 1941, with an average sulphur content of 42 per cent. The average value f.o.b. mines was $3.09 per long ton. Spanish pyrite, 48 per cent sulphur minimum, was quoted at 12 cents a unit c.i.f. U. S. ports, equal to $5.76 a ton for shipments of minimum sulphur content. Most of the pyrite produced in the United States is obtained as a by-product in the dressing of copper and zinc ores. Metal mines in the East making pyrite concentrate for the market or use in company-operated sulphuric-acid plants

Jan 1, 1944

By David R. Mitchell

The mineral pyrite (or marcasite) occurs in coal beds as balls, lenses, veinlets and bands. Several million tons are wasted annually on the refuse dumps from coal mining and coal-preparation activities. Sporadic attempts have been made in the United States to recover this pyrite for manufacture of sulphuric acid but only a few plants have succeeded in recovering it on a commercial scale. Several chemical companies in the East became concerned about supplies of pyrite early in 1943, but by late summer of that year the situation had eased. If for no other reason than national security and self-sufficiency during wartime, methods of recovering pyrite from coal-mine refuse should be worked out and made available to coal-mine operators who are potential producers of pyrite. Eastern sulphuric-acid plants are particularly likely to have their sources of raw material blocked by war conditions, so that special studies to determine where supplies of pyrite can be obtained for these plants from inland sources are desirable. Davis&apos;s1 summarization of conditions during World War I gives the following facts: During those years, the shortage of pyrite became acute. Shipments of hand-picked pyrite were made from a number of coal mines in Illinois, Indiana, western Kentucky, Ohio and Pennsylvania. A mechanical concentrating plant for sepa- rating pyrite from picking-table refuse was built at Danville, Ill., and operated for several years, and a small pyrite-concen-trating plant treating washery refuse near Gillespie, Ill., operated for several months in 1918. The Consolidation Coal Co. of Fairmont, W. Va., constructed and operated a sulphuric-acid plant using pyrite recovered from coal-mine refuse. The U. S. Bureau of Mines during 1917-1919 made a comprehensive survey of pyrite-production possibilities from coal mines and estimated that there was a potential yearly production of 435.000 tons from Ohio and Pennsylvania and 928,000 tons from Kansas, Missouri, Iowa, Illinois, and Indiana mines. This estimate included only mines that could produce one per cent or more of their coal production as pyrite containing a minimum of 40 per cent sulphur. PyRite Production According to the Minerals Yearbook,2 approximately 660,000 long tons of pyrite was produced in the United States in 1941, with an average sulphur content of 42 per cent. The average value f.o.b. mines was $3.09 per long ton. Spanish pyrite, 48 per cent sulphur minimum, was quoted at 12 cents a unit c.i.f. U. S. ports, equal to $5.76 a ton for shipments of minimum sulphur content. Most of the pyrite produced in the United States is obtained as a by-product in the dressing of copper and zinc ores. Metal mines in the East making pyrite concentrate for the market or use in company-operated sulphuric-acid plants

Jan 1, 1944

By J. J. Lillie

Sublevel stoping was first developed in the Michigan iron mines many years ago. Since that time this method, and modifications with long hole drilling, have been used in a number of non-ferrous mines and have been described in various papers and articles. With a few exceptions, the operations where sublevel methods have been applied are producers of large daily tonnages from ore bodies of considerable size. In selecting mining methods for small properties or relatively small detached ore bodies, sublevel methods often have been overlooked. The method has been used successfully in relatively small operations at Copper Canyon, Nevada, and, more recently, in mining small ore bodies at Darwin, Calif. As a result of this experience, we believe that a review of the general procedure, the advantages noted, and the results obtained will be of interest to many operators of relatively small mines. General Description Briefly, sublevel stoping consists of, first, driving sublevel drifts at predetermined vertical intervals through the length of the stope block; second, breaking an initial mining slot across the stope; third, driving crosscut benches from each sublevel by slabbing into the slot; and, fourth, breaking the ground between benches. Applicability Since this is an open stope method, it requires wall rock that will stand over reasonable stope lengths without caving or excessive slabbing. Ore must be moderately firm but need not be hard. The method is best adapted to steep dipping, tabular or lenticular ore bodies but may be applied to ore bodies of almost any shape if their vertical extent is appreciable. It can be used with dips that are too flat for successful shrinkage stopes and can be modified for use with slushers on dips below about 40". Initial mining at the Copper Canyon operation consisted of sublevel stopes for the wider ore sections and shrinkage stopes for the narrow sections where ore widths averaged about 15 ft. After some experience and a study of comparative results on completed stope blocks, sublevels were substituted for shrinkage stopes even in the narrow ore widths. Practical stope lengths depend upon the character of the walls and the continuity of values within the ore body. Where blocks of low grade or waste are included in the stoping limits, these areas can readily be left as pillars for intermediate wall support and longer stope lengths may be planned. Raises can be spotted in narrow or low grade areas and stope lengths adjusted accordingly. Stoping Details To illustrate stoping details, assume that we have an ore shoot 15 to 30 ft wide and about 200 ft long, with walls that would stand over a stope length of 80 to 90 ft (Fig1). PRELIMINARY DEVELOPMENT Preliminary development for sublevel stoping would be simple. A service raise is driven at about the center of the block and stope limits planned to leave a pillar around this raise. Sublevel drifts driven each way from this central raise will outline the

Jan 1, 1950

By E. C. Dominguez, H. B. Scharf

THE Sparrows Point plant of Bethlehem Steel Co., southeast of Baltimore, is the largest Bethlehem plant, with an annual rated capacity of 6.2 million ingot tons. There is considerable diversity of product, but a high percentage is sheet, tin plate, galvanized sheets, and rod and wire. The large amount of light-weight end product requires a number of pickling installations that necessarily consume large quantities of sulfuric acid. In the coking operations sulfuric acid is also consumed in recovery of ammonia as ammonium sulfate and in treatment of recovered light oil derivatives. Total sulfuric acid consumption in the plant is currently approaching 10,000 tons per month, of which 55 to 60 pct is required for pickling. In 1951 the acute sulfur shortage caused grave concern as to future sulfuric acid supply for the Sparrows Point plant. Raw materials for acid making were already at hand in the form of hydrogen sulfide and pyrites sufficient to produce nearly half the sulfuric acid required at that time. It was also known that the Bethlehem mining activities would produce additional pyrites in the future. Late in 1951 Bethlehem management decided to build a 250-ton sulfuric acid plant at Sparrows Point to utilize these raw materials. The plant was built and put in operation in May 1953. The steel plant has expanded considerably since 1951, and it is now necessary to operate the acid plant far over rated capacity in order to supply plant demand. Production of acid in 1954 amounted to 93,000 tons. Raw Materials for Acid Making: Coking operations at Sparrows Point produce a residual fuel gas in large quantity which has a net heating value of 525 to 545 Btu. Prior to 1950, plant surplus coke oven gas was sold to the local utility company. In

Jan 1, 1957

By J. H. Jackson, R. B. Fischer

VERY little is known about the properties of relatively pure refractory metals in the cast state since these metals are customarily made by powder-metallurgy methods. Recently, the development of the vacuum arc-melting furnace has made possible the study of large, sound ingots of cast molybdenum. The ductility of molybdenum, at room temperature, may vary considerably. Often the brittle state depreciates the utility of molybdenum, even for the material prepared by powder-metallurgy techniques. At the outset of this program, it was felt that if the causes of brittleness were to be determined it might be possible to reduce the difficulties met in the working of molybdenum. Also, the problem of producing strong, fusion weldrnents of molybdenum might be solved. The conditions required to produce ductile molybdenum by the powder-metallurgy technique are reasonably well known. A fibrous or elongated-grain structure has good ductility although this property is probably directional. The equiaxed structure obtained in fully recrystallized molybdenum has been considered to be brittle. The reasons for this behavior of molybdenum have not been established. The preparation of cast molybdenum ingots by the vacuum-arc method was described by Parke and Ham.&apos; By this method, ingots weighing up to 250 Ib were produced." Parke and Ham1 found that molybdenum ingots containing more than approximately 0.0025 pct oxygen could not be hot forged. On the other hand, they found that fully deoxidized cast molybdenum, containing up to 0.06 pct carbon, could be hot forged. However, the as-cast ingots containing either carbon or oxygen were brittle at room temperature. Parke and Ham,1 Woodside,3 and Zapffe, Landgraf, and Worden4 noted the inter-granular precipitation of oxide or carbide in cast molybdenum. This study—part of an extensive program on mo-lybdenum—had the following objectives: (1) To investigate the causes for the brittleness of cast molybdenum at room temperature, (2) to investigate the means for improving the ductility of cast molybdenum at room temperature, and (3) to add to the knowledge of cast molybdenum and of cast metals in general.

Jan 1, 1951