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By L. A. Wright, H. L. McKinley

PINE Creek mine of Union Carbide Nuclear Co. is some 23 miles northwest of Bishop, Calif., in the Sierra Nevada Mts. Office and mill are 7800 ft above sea level, the 1500 level portal is at 9300 ft, the A level portal at 10,800 ft, and the 3050 portal at 11,500 ft. Original claims were located in 1916, and the mine worked intermittently until 1936, when it was purchased by Union Carbide Corp. Since then the property has been operated continuously, except for two brief periods when operations were suspended because of manpower shortages. Mine production through 1956 amounted to 2,792,425 dry tons of ore.

Jan 10, 1957

By Bartlett W. Paulding

Since its introduction to the petroleum industry in 1949 by Clark,1 hydraulic fracturing or "hydrafracing" has developed into a widely respected and useful technique for increasing production from petroleum reservoirs. Even then, Clark described production increases of from 6 to 100 barrels of oil per day (bopd). Since that time, others have described treatments in which production increased from 12 to 456 bopd, from only a show to 40 bopd;2 from 48 to 219 bopd; 3 from 500 to 1400 bopd and from a show to 192 bopd.4 Although representative of only a few of many successful hydraulic fracturing operational6 these increases in production attest to the past and potential value of the hydrafracing technique to the petroleum industry. The technique has received a considerable amount of theoretical and experimental attention.1, 5-20 Unfortunately, however, the degree of general understanding of the mechanics of hydraulic fracturing is still inadequate to solve many recurring problems. The practicing engineer is plagued with difficulties in predicting such factors as: (1) the required pressure and viscosity of the fracturing fluid to not only induce the fracture but to control the extent of its growth, (2) the strength characteristics and amount of proppant to include in the fluid so that the fracture is held open once the fluid pressure is decreased, and (3) the volume of the reservoir influenced by the induced fractures and the resulting increase in production. A major cause of these difficulties is the uncertainty associated with determining the orientation of the hydraulically induced fractures. One's natural conclusion in considering the problem is that the orientation of the fracture would depend on several variables, as, for example, the relationship between the viscosity and injection rate of the fluid and the

Jan 1, 1968

By Bernard Moore

AT present the demands of the United States for strontium are met by imports from Germany, England and Canada, which vary considerably in proportions of ore and finished salts, in tonnage and in value. Statistics for the period 1926-1932 are given in Table 1. Domestic deposits are TABLE 1.-Strontium Imported into the United States, 1926-1932a Ore Chemicals Year Short Tons Value Pounds Value 1926 2,200 $58,159 5,433,044 $115,436 1927 1,356 58,711 4,233,816 149,383 1928 759 31,157 3,380,464 136,561 1929 973 45,505 4,578,327 196,233 1930 220 10,459 2,153,788 104.807 1931 (260 lbs.) 123 1,742,714 83,564 1932 5 276 469,481 20,910 a Mineral Industry (1934) 42, 61. known in many states, but the, larger and purer ones are in the West, particularly in the desert region of southeastern California and western Arizona. These deposits are large enough to command interest, par-ticularly in consideration of future cheap power at the Boulder Dam. The deposits described in this paper were examined during an investi-gation of the mineral resources of the Boulder Dam region by the United States Geological Survey for the United States Bureau of Reclamation in the spring of 1934. Mr. D. F. Hewett, who had general charge of the investigation, Mr. W. T. Schaller and Mr. M. I. Goldman contributed many helpful suggestions. Mr. N. H. Harshman assisted efficiently during the field work.

Jan 1, 1935

By Fred C. Bond, Walter Maxson

Since the last publication of tabulated results of grindability tests by the authors1 the total number of ball-mill tests made has more than doubled, and rod-mill tests have become increasingly important. Nearly all of the standard closed-circuit ball-mill and rod-mill tests made to date are included in the present tabulation, together with additional information not previously published. Standard Ball-mill Grindability Tests The method of conducting these closed-circuit tests has not been altered. A sample of the ore or other material is stage-crushed in rolls set at 1/16-in. opening with a 6-mesh screen until all of it has passed the screen. The combined minus 6-mesh screen undersize is mixed, sampled and screen-analyzed, and its apparent specific gravity is determined by packing and shaking in a standard container, and weighing. The apparent specific gravity ordinarily is about 60 per cent of the true specific gravity. The unit volume present in the mill .in all tests is 700 C.C. of the packed minus 6-mesh roll product, and the number of grams occupying 700 C.C. is the unit test weight. This weight is placed in the mill dry, ground for the number of revolutions estimated to be necessary, discharged, and screened mechanically in three testing sieves of the mesh size at which the test is to be conducted. The oversize, or circulating load, is weighed, sufficient fresh minus 6-mesh feed is added to bring the total weight up to that of 700 c.c., and the charge is returned to the mill for a second grinding period. All standard ball-mill tests are conducted at 250 per cent circulating load, and the number of revolutions in the mill necessary to obtain this circulating load at any grinding period is estimated from the results of the previous period. The number of net grams of screen undersize produced per revolution of the mill approaches its final equilibrium value after several grinding periods, and this is recorded as the grindability, or relative ease of grinding, of the sample. A cylindrical ball mill, 12 by 12 in. inside, with a smooth lining and rounded corners, is used with a revolution counter Under standard grinding conditions it runs at 70 r.p.m., and contains a charge of 285 iron balls weighing 20,125 grams, ranging from I½ to ¾ in. in diameter. Tests are conducted at the mesh size to which the ore is to be ground in practice, and the capacity- to be expected from a given mill is calculated by comparing the grindability with that of a similar standard ore whose performance characteristics arc known. No satisfactory general formula for convertillg standard grindabilities into corn; rnercial power-consumption values has been

Jan 1, 1943

By Walter Maxson, Fred C. Bond

Since the last publication of tabulated results of grindability tests by the authors1 the total number of ball-mill tests made has more than doubled, and rod-mill tests have become increasingly important. Nearly all of the standard closed-circuit ball-mill and rod-mill tests made to date are included in the present tabulation, together with additional information not previously published. Standard Ball-mill Grindability Tests The method of conducting these closed-circuit tests has not been altered. A sample of the ore or other material is stage-crushed in rolls set at 1/16-in. opening with a 6-mesh screen until all of it has passed the screen. The combined minus 6-mesh screen undersize is mixed, sampled and screen-analyzed, and its apparent specific gravity is determined by packing and shaking in a standard container, and weighing. The apparent specific gravity ordinarily is about 60 per cent of the true specific gravity. The unit volume present in the mill .in all tests is 700 C.C. of the packed minus 6-mesh roll product, and the number of grams occupying 700 C.C. is the unit test weight. This weight is placed in the mill dry, ground for the number of revolutions estimated to be necessary, discharged, and screened mechanically in three testing sieves of the mesh size at which the test is to be conducted. The oversize, or circulating load, is weighed, sufficient fresh minus 6-mesh feed is added to bring the total weight up to that of 700 c.c., and the charge is returned to the mill for a second grinding period. All standard ball-mill tests are conducted at 250 per cent circulating load, and the number of revolutions in the mill necessary to obtain this circulating load at any grinding period is estimated from the results of the previous period. The number of net grams of screen undersize produced per revolution of the mill approaches its final equilibrium value after several grinding periods, and this is recorded as the grindability, or relative ease of grinding, of the sample. A cylindrical ball mill, 12 by 12 in. inside, with a smooth lining and rounded corners, is used with a revolution counter Under standard grinding conditions it runs at 70 r.p.m., and contains a charge of 285 iron balls weighing 20,125 grams, ranging from I½ to ¾ in. in diameter. Tests are conducted at the mesh size to which the ore is to be ground in practice, and the capacity- to be expected from a given mill is calculated by comparing the grindability with that of a similar standard ore whose performance characteristics arc known. No satisfactory general formula for convertillg standard grindabilities into corn; rnercial power-consumption values has been

Jan 1, 1943

By T. W. Lippert

In these opening remarks, I will endeavor to give a background of continuous casting of metals. Only passing attention will be given to the different processes to be described by the authors whose papers are on the program, as they are far more familiar than I am with their subject matter. Suffice to say that over the past several years the brass, copper, aluminum and magnesium industries have shifted practically completely from conventional static molds to continuous casting machines. The Aluminum Company of America alone has more than 60 continuous machines in operation. The product as continuously cast is superior metallurgically, less scrap loss is involved, costs are generally lower, and the equipment requires little floor space. A continuous machine, operating as it does with deadly constancy, is a ravenous consumer of molten metal, and in most plants one or two machines are well able to handle large banks of melting and refining furnaces. In steel, interest in continuous casting is today more serious than ever before in history, and there is every reason to believe that continuous machines may secure a minor place in the ferrous industry, particularly for tool and alloy analyses. In the past, steel has been commercially cast continuously, and today many tons of steel are being turned out in this manner. It has been said that the cradle of every science is surrounded by dead theorists— and certainly there are innumerable examples of this in the metals industries. An excellent illustration, for instance, is the continuous rolling of metal, a tantalizing objective that for decades on end exhausted . the efforts of a long line of famous, infamous and unknown inventors. Finally, in 1922, John B. Tytus, of the American Rolling Mill Co., brought forth what now seems such an obviously simple scheme— a concave roll face, each stand of progressivcly less concavity. And out of this one change in design sprang today's wide continuous mill rolling steel sheet Ioo in. wide at 2000 ft. per minute. Another excellent example is the con- . tinuous casting of liquid metal into an intermediate form, preferably one requiring little subsequent breakdown. First, the casting of liquid metal alone was fascinating enough. Then someone passed the cast metal directly into a rolling mill, and shortly thereafter this continuously rolled strip was passed directly into a stamping machine. This leitmotiv of a continuous flow of metal from the liquid phase to a consumer end product, fast and inexpensive, and all housed within a small space, was enough to drive some 200 or so inventors into frenzy over the past hundred years. All the ideas, however, like mayflies, had only their brief afternoon of existence. They were born and they died on paper, as constant exemplification of that great void that always exists between mental production and commercial production.

Jan 1, 1945

By Edmund D. North

In making a glass model of mine-workings, each mine will present some little individualities, to meet which will call for the exercise of special ingenuity. Having made several models, I offer the following details of the model of the Montana-Tonopah mine-workings, believing that with small changes for different conditions the method described can be successfully used in making a glass model of the workings of any mine. The model, Fig. 1, has the top, one end, and the front removed. The case is made of kiln-dried pine, painted with several coats of dead-white paint. It is made to take sheets of glass 24 by 42 in. in size, and is 18 in. high, the workings shown being on a scale of 1 in. = 40 ft., or 4800 : 1 scale. The base is made of four pieces of 2-in. material, well pinned and glued together. The back, of material 1 in. thick, is secured rigidly to the base by screws and countersinking. The four posts, 1.5 in. square, are so cut that the front and two ends of the model fit in flush with the outside edges of the posts. The top rails, 1 in. square, are fastened rigidly to the front and back posts by means of small iron braces. The ends, front, and top of the case are of 0.5-in. material; the ends and front are held in place at the bottom by pegs fitting into the base ; at the top the ends are held by buttons and the front by cupboard-hooks on the inside. Hinged hasps on the cover, fitting over staples in the end-pieces, make it easy to lock the entire case. Before the back is fastened to the base, it is cut with horizontal and vertical grooves to receive the horizontal and vertical sheets of glass, shown clearly in Figs. 1 and 2. The horizontal grooves, spaced to scale, represent the vertical distances between the levels at the shaft. The vertical grooves are placed wherever cross-sections of the workings are desired.

Jan 1, 1910

By Arthur Taggart

INTRODUCTION THE flotation process for the concentration of ores is a method by means of which one or more of the minerals in the ore (usually the valuable ones) are picked up by means of a liquid film and floated at the surface of .a mass of fluid pulp. Here they are separated from the other minerals, which remain immersed in the body of the pulp. In general, the minerals which are floated are sulphides of metallic luster, but some other minerals of metallic luster such as graphite and some sulphides with adamantine luster, such as sphalerite and cinnabar, are amenable to treatment by the process. The importance of flotation lie's in the fact that it is primarily a "slimes process" by means of which the particles of valuable mineral, too fine for efficient gravity concentration, are saved with a high percentage of recovery. Recoveries in the mills treating low-grade copper sulphide ores have been advanced 10 to 20 per cent. by the installation of the process and similar increased savings have been accomplished by the same means in mills treating sulphide ores of zinc and lead. When finely ground ore containing sulphides mixed with a siliceous or earthy gangue is brought gently onto the surface of a body of water, in a direction forming an acute angle with the surface of the water, a considerable portion of the sulphide constituent of the ore floats on the surface of the liquid, while the gangue sinks. This is the so-called film flotation, exemplified by the Wood and Macquisten processes. When gas bubbles are introduced into a fluid pulp composed of finely ground ore and water, to which has been added (1) a small amount of certain oils, or (2) a small amount of certain acids or acid salts, or (3) a small amount of certain alkalis or alkaline salts, or (4) a small amount of a mixture of oil with acid or alkali, the sulphide particles in the ore are brought to the surface on the gas bubbles. These collect in a froth

Jan 8, 1916

By Louis A. Panek

Caving of previously undisturbed ground was conducted for a period of about 20 months at two locations about 600 m (2000 ft) apart in an Arizona porphyry copper deposit. When mining was suspended the width of the mined zone at both sites was on the order of one-half the depth and the thickness of ore extracted was only a fraction of the rock column. Thus the results have potential application to predicting the subsidence above a cavern-type excavation. Break cables were used to detect the upward progress of caving. Analytical aerotriangulation was used to determine the vertical and horizontal displacements of 400-plus surface targets covering a 1585 m (5200 ft) square area. The measurements are summarized and interpreted with a view to predicting the rate of rise of a caving brow, detecting the initial surface subsidence, and characterizing the extent, the location, and the varying degrees of expected damage to surface facilities.

Jan 1, 1984

By Dale K. Fields

With today's technological advances in mineral processing equipment, there is an increasing demand for accurate system control. This can only come about through representative sampling. The purpose of this paper is to guide in the selection of sampling equipment and outline design parameters for the cutters used in dry and wet sampling without adding redundant mathematical probability formulas. Special attention will also be given to the equipment used in conjunction with samplers and sampling system, such as feeders, crushers, chutes, launders, and controls.

Jan 10, 1979

By W. A. Schaeffer, M. Albertson

The Yates field is situated in northeastern Pecos County, Texas. Its productive area is about 15,000 acres. Production is chiefly from limestone of Permian age. Depths of wells vary from 1100 to 1800 ft. Initial productions range from a few barrels to 8000 bbl. per hour. The field outlet has always been prorated. Gas-oil ratios originally varied from less than 100 cu. ft. up to 8000 cu. ft. per barrel at allowable production rates. The average ratio was about 500 cu. ft. Fig. 1 shows the structure of the field and the location of the Tippett lease, on which the original experimental tubing work was done. Before the latter part of the year 1929, little actual effort in practice had been given to determination of the conditions under which the highest efficiencies may be realized in naturally flowing wells. For the preceding few years attention had been given to gas-oil ratios and their possible reduction by maintenance of back-pressure by choking at the well head. Some attempts had been made toward reduction of ratio by bottom-hole choking. In late 1929, the Shell Petroleum Corp. started the experimental installation of specially designed flow columns adapted to high-pressure flowing conditions in the Yates field. The objectives in view were the attainment of low gas-oil ratios at low rates of flow and the maintenance of high bottom-hole pressures. All naturally flowing wells should be provided with a suitable eductor string immediately upon completion of the well, so that the benefits of efficient flow will be derived from the beginning. This paper presents a method of tubing for efficient flow and a discussion of some complex factors which influence well behavior. Factors Influencing Gas-oil Ratios Many factors influence the production of gas from naturally flowing wells. The writers believe that in some instances gas control is effected mainly in the reservoir itself, in others in the well bore, and in still others both sets of factors may be significant. Some wells apparently have penetrated productive zones in which practically all of the gas is closely associated with, and mostly dissolved

Jan 1, 1932

By E. S. Bardwell

The reverberatory copper-refining furnaces at the Great Falls Reduction Dept. of the Anaconda Copper Mining Co. have used successively as fuel, lump coal on grates, pulverized coal, oil and natural gas. In no other plant, so far as can be ascertained, has such a variety of fuels been used. This fact affords a unique opportunity for comparing the relative efficiency obtained with these several fuels under nearly identical conditions. Furnaces The present furnace refinery at Great Falls was placed in operation in April, 1916. The original installation consisted of two furnaces, each with hearths 36 by 14 ft. and provided with fireboxes 10 by 12 ft. for firing lump coal on grates. These furnaces had air-cooled sand bottoms and chrome-brick side walls both above and below the metal line. The roofs were of 12-in. silica brick. Each furnace was provided with a waste-heat boiler of approximately 400 boiler-hp. capacity. We soon found that chrome brick was very unsatisfactory as side-wall material because of its excessive spalling. At the first opportunity the furnaces were rebuilt, using magnesite brick below the metal line and silica brick from the metal line to the roof. At the same time the thickness of the side walls was increased, thus decreasing the width of the hearth to 13 ft. At about this time also, the thickness of the furnace roofs was increased from 12 to 15 in. No further changes occurred in furnace design until 1922, when the use of lump coal on grates was abandoned in favor of pulverized coal. The first charge to be melted and refined with pulverized coal was cast on March 30, 1922. When pulverized coal first was used the burner pipes were introduced directly into the firebox as it was felt that some kind of a combustion chamber might prove necessary, but this was not needed. The fireboxes were dismantled and the burners introduced through openings directly over the bridge wall. During the summer of 1922 one of the furnaces was completely rebuilt, lengthened 10 ft. and provided with a silica-brick bottom in the form of an inverted arch. At

Jan 1, 1934

By J. D. Miller, R. J. Smith

The hot water process for Utah tar sands differs significantly from that used for Canadian tar sands due to inherent differences in respective bitumen viscosities and the nature of bitumen-sand association. These differences have led to the identification of certain design criteria necessary to achieve satisfactory phase disengagement during digestion (Sepulveda and Miller, 1978; Misra and Miller, 1980). Subsequent flotation of digested tar sand is a subject of current attention and is considered in this publication for the Asphalt Ridge deposit. Although contact angle measurements of solvent extracted Asphalt Ridge bitumen indicated moderate hydrophobicity, air bubble attachment to the bitumen concentrate is not possible. This surprising result suggests that flotation separation is dependent on air bubble entrapment rather than on air bubble attachment due to surface hydrophobicity. The entrapment of air bubbles in the bitumen was apparent from examination of the bitumen concentrate, especially at lower flotation temperatures. A factorial design of the major operating variables in flotation separation indicated that the quality of separation was significantly dependent on the flotation temperature and to a lesser extent on the degree of agitation. For a flotation temperature of 77°C a recovery of 96.7% was realized at a grade of 61.0% bitumen. The improved separation at higher flotation temperatures was found to be due to the decrease in bitumen viscosity resulting in more effective rejection of coarse sand from the concentrate. This research effort, which has been sustained for the past six years at the University of Utah, is being continued to provide necessary design data for the construction of a pilot plant to treat Utah tar sands.

Jan 1, 1982

By H. R. Hanley, Charles Y. Clayton

ONE of the most important contributions to the art of electrolytic zinc production has been the aluminum cathode. This has been used in all major production since its commercial development in 1916. The literature is replete with important contributions, some of which may have been replaced by better ones, but thus far there appears no other metal to take the place of aluminum as cathode material. Credit for this innovation goes to Sherard Cowper Coles, of England. When experi¬menting on the extraction of zinc from Broken Hill ore at Hayle, Cornwall, in 1893, he introduced the aluminum cathode, and in 1895 obtained British Patent No. 20073. The reason for this dominant position of aluminum as cathode material in zinc sulphate electrolytes is its unique surface, to which electrodeposited metals usually do not tightly adhere. The metal must be of high purity for best results. It is possible to develop a starting-sheet method by which thin electrodeposited sheets of cathode zinc are made and stripped from a smooth aluminum plate. Such thin sheets of cathode zinc may be used as are the copper sheets in standard electrolytic copper-refinery practice. However, complications arise because of extreme warping of the zinc starting sheet after a few hours of deposition, and thus special arrangements must be made to compensate for this abnormality. Reasons advanced for the ease with which deposits may be stripped from the aluminum surface are: (1) presence of a permanent oxide coating on the metal, (2) failure of aluminum to alloy with the deposit, (3) differences in coefficient of expansion, (4) high position of aluminum in the electromotive series of metals. Of these reasons for nonadherence, which may have some application in general deposition, probably the oxide coating has the greatest bearing on the subject of zinc deposition. Adherence of zinc deposits has occasionally had serious effect oil tonnage and production costs. Opinions have differed as to the specific

Jan 1, 1938

By Oscar A. Glaeser

The purpose of this paper is to present specific applications of the fundamental principles of mine ventilation and the means that are at our disposal to accomplish the task of maintaining a healthful atmosphere in small mines and prospect openings. Need for Ventilation The range between the maximum oxygen content of normal air and the minimum safe content is extremely narrow, approximately 5 per cent. Furthermore, normal air must not be contaminated with irrespirable gases beyond definite limits, and mineral dusts are equally dangerous, although health impairment is slower. Tables 1 and 2 present the more common respiratory hazards encountered in mines. Adequate ventilation is the most effective means available to reduce or eliminate these hazards. Mechanical ventilation has reached varying stages of perfection in coal mines and in many large metal mines. The small mines and prospects rarely have a definite plan of ventilation. As a consequence underground workers are frequently exposed to health and accident hazards. The annual reports of fatalities in mines in the Western States record yearly many deaths due to asphyxiation and deaths caused by men falling down shafts and raises after being overcome by gas. At times mining operations are slowed up because there is no means of driving out the gas after blasting. For that reason only one shift is worked, or men wait from one to several hours for the "smoke" to clear out. Methane gas in varying quantities is widely distributed throughout California. It has been encountered in explosive concentrations in the coast range. It has been found on the Mother Lode and in quicksilver mines near the coast. Methane explosions have not been uncommon in underground operations and in tunneling jobs. These conditions and records indicate that there is room for improvement in mine-air conditions.

Jan 1, 1937

By Stanley F. French

ELECTRICAL power and distribution for the sulphide ore plant is an addition and extension of the existing power system which has served the oxide reduction plant since its inception in 1914. The power plant is located at Tocopilla on the west coast of Chile, some 90 miles from Chuquicamata. A brief description of this plant and the existing transmission system is necessary to an understanding of the problems involved in the provision for the sulphide plant load. Originally the Tocopilla power plant consisted of four Escher Wyss turbines driving Siemens Schuckert generators, each rated 10,000 kw. At various times, units of 10,000 and 20,000 kw were added, all of which operated at 215 psi and 600°F at throttle. In order to improve the economy of the station, a Topping unit was installed in 1938 together with a high-pressure steam generator of 360,000 lb per hr capacity. This unit is rated at 11,000 kw and was designed for 750 psi, 880°F. The exhaust from this turbine passes to the low pressure station header for operation of part of the low pressure turbines. Total kw generated by the high pressure boiler, topping turbine and associated low pressure turbines totals 33,000 kw and the total station capacity is 113,000 kw. This installation proved one of the most successful of its kind, operating almost continuously. The original plant was of German design and consequently European characteristics were incorporated into the system, the frequency being 50 cycles and voltages of 500, 5000 and 100,000 were used for low, generator, and transmission voltages,

Jan 1, 1952

By Robert V. Whitman

INTRODUCTION Ideally, this paper should start by citing examples of slope failures in open pit mines during earthquakes. I do not know of any such examples, although - since my search of the literature has not been exhaustive - such examples may well exist. However, there certainly are many examples of devastating earthquake-induced landslides, and there is no reason why such landslides cannot occur at large open pits. Moreover, many of the actual earthquake-induced landslides have seriously affected mining operations in one way or another. Hence, this paper will treat earthquake-induced landslide phenomena in general, in the hope and expectation that the reader can apply this general knowledge to the specific problems associated with open pit mining. Ultimately, of course, the aim is to predict the likelihood of the occurrence of earthquake-induced landslides, and to delineate steps which can be taken to minimize the possibility of their occurrence. These aims can be achieved only imperfectly at the present time, for our knowledge concerning earthquake-induced landslides is still in an embryonic stage. Nonetheless, much useful knowledge does exist. In order to understand and apply this knowledge, it is necessary to: 1. Review the nature and extent of actual earthquake-induced landslides; 2. Have a grasp of the basic terminology and concepts of seismology; 3. Study the fundamentals of probability-based seismic risk analysis; 4. Appreciate the way in which earth materials behave during earthquake loadings; and 5. Have a general knowledge of the methods currently available for analyzing the dynamic response of slopes. Much of this paper is devoted to this necessary background. Information of practical use in connection with various classes of slope stability problems is assembled in the final sections. LANDSLIDES DURING PAST EARTHQUAKES Almost every earthquake in mountainous country, strong enough to be reported in the literature, has caused at least some landslides and rockfalls. Unless these slides or falls damage

Jan 1, 1971

By I. Milton LeBaron

OPERATION of the phosphate flotation plant at the Peace Valley mine of the International Minerals and Chemical Corporation, Florida, began in the summer of [ ] 1942. Since that time several changes have been made in the flowsheet, with resultant increases in metallurgical results. Three flowsheets, together with the comparative plant results, are given in this paper, together with a discussion of the desliming circuit used for preliminary washing of the flotation feed. FIRST FLOWSHEET CHANGES Fig. I lists the essential units in the original flowsheet at the Peace Valley plant. The figures for tonnage and grade are average for a period of two months. Mineral Separation Airflow cells are used for both the fatty acid and the amine float. Three immediate difficulties presented themselves in the use of this flow sheet: I. Floating a maximum amount of phosphate in the rougher section to give good recovery sent a low-grade concentrate to the silica float, resulting in high costs for amine reagent in that section.

Jan 1, 1946

By A. R. Kaufman, P. Gordon, C. H. Schramm

AS a part of the general program on alloys of uranium carried out at the Massachusetts Institute of Technology under contract W-7405-eng-175 for the Manhattan Project during the recent war, it was considered desirable to investigate the boundary binary phase diagrams of the ternary alloy system uranium-tungsten-tantalum. Previous work on these alloys was practically nonexistent. A brief statement to the effect that tantalum and tungsten form a continuous series of solid solutions was found in a general paper on tantalum by W. V. Bolton1 but no information was available on the uranium-tungsten and uranium-tantalum systems. The research to be described in this paper has served to delineate the main features of the uranium-tungsten, uranium-tantalum phase diagrams and to check Bolton's work on the tungsten-tantalum diagram. Experimental Details Metals: The tungsten and tantalum used for preparing the alloys were 99.96 and 99.65 pct pure, respectively. The chief impurities in the metals were about 0.02 pct molybdenum in the tungsten, and 0.2 pct columbium, 0.05 pct tungsten, 0.02 pct barium and 0.01 pct silicon in the tantalum. The uranium used was 99.9 pct pure, with the main contaminants being iron and carbon. (An appreciable amount of magnesium was present in the uranium as received, but this distilled off during melting and so is not reported as an impurity.) The composition of specimens indicated in the subsequent figures and tables of this paper are the results of chemical analyses on the cast alloys except for the tantalum-tungsten alloys. For the latter, chemical analysis was not very reliable as a result of the difficulty of separating tungsten and tantalum chemically. Consequently, the nominal

Jan 1, 1951

By James P. Kimball

THE object of the present memoir is to put on record some practical experiments by the writer in smelting a silicious native magnetite with no other flux than the silicates of its own gangue: This work was undertaken in the spring of 1878 for Messrs. Naylor & Co., of Boston, to whom I am indebted for the privilege of here communicating its results, these having hitherto only been simply announced.* So far as I am aware these are the first instances of the kind in this country that non-calcareous; or rather non-carbonated, magnetites have successfully been used strictly alone as self-fluxing ores in the blast furnace; that is, without the addition of calcareous flux in the form either of limestone, quicklime, calcareous ore, or basic (calcareous). slag. Previous to the spring of 1878, when the Chateaugay magnetic ore was first brought to my notice, its use had been limited to Catalan forges, after "separation" at the ore bed. The blooms and billets made from it in the several forges of Messrs. Weed & Williams, and of Mr. Williams, were well known in the trade as among the best in the market. The Chateaugay ore beds proving of great magnitude, and the ore exceptionally low in phosphoric acid, it commended itself for blast-furnace purposes to Messrs. Naylor & Co., who employed nee to blow-in the charcoal furnace at Pittsford, Vt., on this ore in the spring of 1878., a previous trial of this ore in the same furnace in the fall of 1877 having proved abortive. This was therefore the first time that Chateaugay ore was successfully smelted in a blast furnace, although since well known as a blast-furnace ore possessing superior and exceptional qualities. The Pittsford furnace, although rebuilt in 1844, and subsequently improved by the former owner, the late J. Prichard, is one of the few furnaces of the old type with open breast (Sumpfofen) still active in this country. Its improvements on furnaces of the period, * Bulletin American Iron and Steel Association, XII, 1878, 132, 156.

Jan 1, 1881