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By E. E. Carpenter

Believing that the results accomplished in the mill of the Churchill Milling Co., Wonder, Nev., during the fiscal year ended Sept. 30, 1914, will be of interest, I am presenting the more prominent facts and figures showing the progress made in treating this ore with especial reference to tonnage, supplies, costs, and extractions. The mill is situated at Wonder, Nev., in the Wonder mining district, 55 miles east from Fallon, Nev., the nearest railroad point. All supplies are hauled this distance by freight teams, this method of transportation having been found more reliable for continuous service than truclis, the price paid for hauling being slightly under one cent per pound. In this mill are treated the ores of the Nevada Wonder Mining Co. Ore is Unusually Hard The company which tested this ore before the mill was built reported "This ore is an unusually hard, tough, oxidized quartz, carrying a small percentage of high-grade sulphides, containing good values in both gold and silver. None of the rebellious silver minerals appear to occur in appreciable quantities." ... "In regard to the mill tonnage this ore is unusually hard and tough, and a high stamp duty cannot be expected. Using a 20-mesh screen and 1,050-lb. stamps, our best work was three tons per stamp per 24 hours." It is very probable that the ore delivered to the testing plant was harder than the regular mill run has proved to be. However, at present the mine is developed to a depth of 1,000 ft., and still shows hard ore in places at this depth in about the same proportion as on the upper levels from which the test ore came; during this year's work, ore was treated from nearly all levels. Together with this hard ore in the mill feed a considerable proportion appears to be broken down and tending to crumble, thus making an ideal mixture for stamp milling. A small proportion of talc is also noticeable, which sometimes increases in amount to the point where it is very troublesome in the thickeners.

Jan 1, 1916

Capable of processing 3500 tpd of uranium ore, the United Nuclear-Homestake Partners' mill in Grants, N. M., differs considerably from the nearby mills of Anaconda and Kerr-McGee in at least two major aspects: it leaches uranium ore with an alkaline solution of Na2C03 and NaHC03 instead of sulfuric acid; and it recovers yellowcake directly by precipitation from the pregnant alkaline leach solution, without any need of ion exchange or solvent extraction circuits. A first look at the flowsheet shows that this mill has two separate grinding and leaching circuits. Reason for the double circuitry: the mill treats, on a toll basis, uranium ores for Anaconda as well as for the partnership, and it must also treat ores coming from mines separately owned by Homestake Mining Corp. and from mines owned by United Nuclear Corp.

Jan 8, 1974

By W. Joseph Schlitt

The characteristics of successful heap and dump leach operations are first reviewed in a general sense. Then attention is focussed on solution management as the key operating parameter in leaching. The first part of this study includes a detailed conceptual analysis of fluid flow. This identifies the relationships between the flow parameters and the chemical and physical properties of the rock. Then these relationships are extended to practical systems in order to establish optimum application rates and solution distribution methods for heaps and dumps. The study shows that there is a maximum effective application rate, qmax, that can be used without causing solution short circuiting. This rate is a function of the minimum permeability in the heap or dump. The expression for qmax is qmax (gal/ft2/hr)[~]0.9 k (darcies). Here k is the appropriate permeability coefficient. Using this equation, qmax is found to be on the order of 1 gal/ft2/hr (4 cm/hr) for a typical heap or dump leach operation.

Jan 1, 1984

By K. G. Colman

This text is intended to provide the operator and turn-key contractor with the basic data for sizing vibrating screens to be used in ore crushing plants. The basic capacities, modifying factors, etc. used in the area requirement calculations were developed empirically and modified to their present values based upon years of practical applications. It must be pointed out that screening is much more of an art than an exact science, however, the information presented here should provide for reasonably accurate screen sizing when applied to metallic ores .

Jan 1, 1978

By Torgrim Eftestoel, Leif Olsen, Ove Sandberg

IN the vertical contact Soederberg electrode for aluminum furnaces more or less serious cracks are sometimes formed in the electrode, with harmful effect on furnace operation. The problem of crack formation has been studied on a laboratory scale to obtain a better understanding of this phenomenon. The primary cause of crack formation is the shrinkage of the electrode carbon, which is opposed by the normal thermal expansion of the steel contact studs. The forces thus set up are active in the temperature interval from which the electrode obtains a rigid structure (450" to 500°C) to furnace operational temperature (950°C). A visualization of our reasoning is given in Fig. 1, which shows the different behavior of mild steel and electrode carbon during baking. Research work performed to determine relationships between crack formation and shrinkage will be described. A brief description of the method developed for measurement of shrinkage will be given, as well as a critical analysis of this method. Also, raw material quality and paste composition variables will be related to shrinkage phenomena. Laboratory Test of Crack. Formation To test the tendency toward crack formations of Soederberg electrodes from different pastes, experiments were performed to simulate conditions of actual commercial electrodes, and are described as follows. A cylindrical test electrode is prepared by tamping the electrode paste to be examined in a sheet iron casing, 350 mm diam by 700 mm height. The casing is perforated, with the holes (2 mm diam) 50 mm apart, to facilitate the escape of volatile matter during baking. A cylindrical steel stud of actual service size, 100 mm diam, is welded eccentrically to the bottom steel plate, and protrudes up through the paste. During baking, a crack may thus occur in the weaker part of the test electrode. The electrode is baked by lowering the assembly slowly into an electric furnace, the temperature of which is kept constant at 900°C. After baking of the electrode, the sheet iron casing is removed and the electrode inspected for cracks. The pictures in Fig. 2 are of three different pastes baked according to the present method. Electrode 1 has a large crack; electrode 2, a somewhat smaller one; and electrode 3, only a minor crack. As will be seen from the simplified sieve analysis diagram of Fig. 2, the compositions of the pastes in question are very different. This subject will be discussed later in greater detail. Crack Formation and Shrinkage Relationships It is well known that shrinkage occurs in coke in the higher temperature ranges of coal

Jan 1, 1958

By Phillip B. Bucky, R. S. Taborelli

The term "longwall mining" is best known to coal men, although modifications of the method are continually being used in other fields. Longwall mining is of interest today because it makes for greater inechanization and better supervision at the working face, and therefore lower production costs and increased safety. It is an approach to the factory system of production, the high efficiency of which as compared to that in mines is well known. There is available an accumulation of experience regarding this work but coordination of experience is lacking and this series of experiments was undertaken to help serve this purpose. Model experiments were used for reasons of safety, cost and time, and because it has been previously shown1,2,3 that a scalar inodel built of the same material as the prototype will behave in a manner similar to that of the prototype, with time effects the same, if the effective weight of the model is increased in the same proportion as its linear dimenbions are decreased. This is accomplished by substituting a centrifugal for the gravitational field; i.e., placing the model in a centrifuge and controlling the strength of the centrifugal field by varying the rate of rotation. Many factors affect the behavior of the geologic material overlying a bed, therefore results and conclusions contained in this paper apply only to the conditions presented herein. Longwall Mining What Is Longwall Mining?—Agreement should be reached as to what constitutes longwall mining and therefore what constitutes a longwall face. Let longwall mining be defined as the mining procedure in which the roof and overlying bed breaks tend to parallel the working face. In the ordinary stope or room-and-pillar mining the roof breaks tend to

Jan 1, 1938

By Phillip B. Bucky, R. S. Taborelli

The term "longwall mining" is best known to coal men, although modifications of the method are continually being used in other fields. Longwall mining is of interest today because it makes for greater inechanization and better supervision at the working face, and therefore lower production costs and increased safety. It is an approach to the factory system of production, the high efficiency of which as compared to that in mines is well known. There is available an accumulation of experience regarding this work but coordination of experience is lacking and this series of experiments was undertaken to help serve this purpose. Model experiments were used for reasons of safety, cost and time, and because it has been previously shown1,2,3 that a scalar inodel built of the same material as the prototype will behave in a manner similar to that of the prototype, with time effects the same, if the effective weight of the model is increased in the same proportion as its linear dimenbions are decreased. This is accomplished by substituting a centrifugal for the gravitational field; i.e., placing the model in a centrifuge and controlling the strength of the centrifugal field by varying the rate of rotation. Many factors affect the behavior of the geologic material overlying a bed, therefore results and conclusions contained in this paper apply only to the conditions presented herein. Longwall Mining What Is Longwall Mining?—Agreement should be reached as to what constitutes longwall mining and therefore what constitutes a longwall face. Let longwall mining be defined as the mining procedure in which the roof and overlying bed breaks tend to parallel the working face. In the ordinary stope or room-and-pillar mining the roof breaks tend to

Jan 1, 1938

By Harmon F. Fisher

In the course of a comprehensive investigation for the development of the electrical process for the dehydration of oil-field emulsions, the writer has had an unusual opportunity to direct and execute a number of related, parallel studies concerning crude oil emulsions and extending over a period of nearly six years. Among these was an investigation utilizing the microscope for studying both the general characteristics of crude oil emulsions, the nature of their electrical charge and the action of an alternating current field in causing coalescence of the finely dispersed, more or less permanently suspended, emulsion particles into larger droplets of such mass that subsequent dehydration would follow by the ordinary process of settling, further agglomeration of the water droplets and, finally, separation into commercially dry, pipe line oil and a clean water discharge. It is not the purpose of this paper to enter into a detailed discussion of the highly specialized subject of microscope technique, nor into a theoretical discussion concerning the origin of the electrical charge on the emulsion particle or concerning the mechanism of the electrical phenomena occurring during the electrical treatment of oil-field emulsions. In a short paper of this character one can hope to touch on only a very limited discussion of the theoretical aspects of crude oil emulsions. The subject of emulsions and colloids is comprised in a comparatively young science in which, however, very rapid progress is being made and consequently it is undergoing many changes; a tentative conclusion valid today may be overthrown tomorrow. Nor would it be a simple science even if it were no longer undergoing this rapid progress; a thorough study of emulsions is based upon a knowledge of such fundamental subjects as molecular structure, ionization and other electrical phenomena, surface chemistry, surface and interfacial tension phenomena, viscosity and other factors too numerous to discuss here and the fundamental theories of which are still undergoing more or less change. Therefore this paper will confine itself largely to a generally descriptive discussion, a sufficiently detailed outline of the microscope technique and a brief comment on

Jan 1, 1931

By Henry J. Schwellenbach

Lack of a system in scheduling the replacement of capital equipment can result in emergency purchases of units which may later be found unsuitable for the job. New York Trap Rock Corp., which produces crushed stone for use as construction aggregate, was faced with just such a situation some years ago. The company replaced capital equipment only when machinery in use was no longer serviceable, or when there was an obvious need for different equipment. Operating and maintenance costs were generally neglected because an efficient method was lacking. New York Trap Rock has four plants located along the Hudson River, all within 70 miles of New York City. Crushed stone output-6 million tons a year-is moved to markets in metropolitan New York, New Jersey, and Long Island areas by water transport. Customers are contractors with ready-mix concrete and asphalt plants located on the water.

Jan 10, 1959

By V. C. Allison

The explosibility of a dust is favored by its fineness and content of volatile combustible matter; its explosibility is decreased by its moisture and ash content. The explosibility of a dust can be reduced by admixture of either inert dust or water. The major effect of both the inert dust and water lies in their heat capacity; they absorb the heat of the explosion flame until the temperature of this flame drops below the ignition temperature of the explosive heat. The amount of inert dust necessary to make an explosive dust non-explosive can be calculated from its proximate analysis. FOREWORD LARGE-SCALE testing of the explosibility of coal dust as conducted by the Bureau of Mines in its Experimental Mine involves a large initial investment, and a high charge for maintenance and conduct of the test; in addition the work is necessarily slow because of the time required for the preparation of the tests and the cleaning and repairing of the mine after each test. These tests have been made along practical mining lines that admit of a range of mixtures of dust varying from 5 to 10 per cent. and with gas as low as 0.5 per cent. increments. Because of the time and expense involved, efforts have been made to devise a type of gallery for laboratory use in which the results obtained in the Experimental Mine may be correlated as an aid in a study of dusts, the explosibility of which might be approximately determined quickly and at a minimum of expense. The explosibility of coal dust has been studied by a number of European investors, using small laboratory devices; among these may be mentioned the experiments con¬ducted by Galloway, Vital, Hall and Clark, Abel, Mallard and Chatelier, Thorpe, Bedson and Widdas, Engler, Holwartz and Meyer, and Taffanel, results of whose experiments have been reviewed by the Bureau of Mines., The Bureau has conducted many tests with the Frazer-Clement2 apparatus, which gives the pressures set up when the dust cloud within is, ignited with a platinum coil or tube, heated to different temperatures, but for these tests the sample must be especially prepared as to size (200 mesh) and must be dry. An attempt has been made to design an apparatus that may be used for checking and correlating the experimental tests. The first design was a wooden gallery, 8 by

Jan 7, 1925

By Jesse O. Betterton

IN the last analysis, two basic factors influence the use of metals and alloys; namely, cost and adaptability to the use under consideration. These are so interrelated that to study the properties of metals or alloys without considering the prices and price trends is quite illogical. Lead has been a cheap metal, selling in 1932 for the extremely low price of 2.56 cents per pound. The cheapness of lead in the past was not due to a particularly abundant supply of the element in the earth's crust but rather to the ease with which it was won from its ores and to the chemical stability of the metal, preventing its dissipation. Today, however, our modern economy has caught up with the distribution of lead in the earth's crust and the price is largely dictated by the supply of concentrates readily available for smelting, and which can be placed quickly upon the market. The demand for lead, particularly in times of stress, has brought the price up to the level of 19 cents in the United States, where the price is controlled by Government decree, and the price has been even higher in the world markets. The higher price for lead has changed the uses for this metal to a marked degree, driving it out of the fields where it is not essential. Lead articles of great antiquity have been found in Egyptian ruins, proving the stability of the element and indicating the basic simplicity of its metallurgy. It is probable that lead was the first

Jan 1, 1953

By G. P. de Forest, G. O. Hiers

The object of the experiments reported in this paper is to determine the eutectic composition of copper and tin and to determine the location of the part of the liquidus line immediately above the eutectic composition in copper. Three methods were used in the investigation: (1) the method of dissolving copper in a bath of molten tin at constant temperature, with the idea that when the bath was saturated with copper its copper content, by analysis, together with the temperature at which the saturation took place would locate a point on the liquidus line; (2) the method of the differential cooling curves, and (3) the method of microscopic examination. This paper is principally concerned with the results of the last two methods mentioned. The results of the first method, that of dissolving copper in molten tin, did not agree with the results of Miller,' by the same method, nor with the results of other investigators by other methods, nor yet with the results of the cooling-curve experiments and microscopic examination reported here. Solution of Copper in Molten Tin To determine the amount of copper that would be dissolved by molten tin at a given constant temperature the apparatus shown in Fig. 1 was used. It consists of a cast-iron pot a, 6 in. deep and 8 in. dia., well cleaned inside and dressed with graphite. This pot, which holds 15 lb. of tin, g, rests on a hot plate b with resistance winding c. The thermocouple d leads to a Wilson-Maeulen temperature regulator for maintaining the temperahre at the heating coils at a fixed temperature to within ±5°C. There is a shallow bath of solder e on the top of the hot plate to improve the thermal contact between the plate and the pot. An insulating jacket f of asbestos prevents the irregular radiation of heat from the apparatus. The copper stirrer h slowly rotates in, and is dissolved by the bath of molten tin g. This bath is protected from excessive oxidation by a layer of charcoal j. The stirrer h is fixed to the lower end of the iron tube k, which is so mounted that it can be driven by a motor with reducing gear. The temperature of the bath is indicated

Jan 1, 1930

By J. N. Ong, J. P. Ratledge, J. H. Boyce

SINCE German operators opened the Tsumeb mine in the early 1900's, continuous operation has been interrupted only by enforced shutdowns during two world wars and the depression of the 1930's. Original metallurgical practice was blast furnace smelting. This was applicable only to low zinc ore, leaving large tonnages of high zinc ore to be stockpiled on the surface. A gravity concentrator using tables, buddles, and jigs operated for several years, but differential flotation of the ore did not get beyond the experimental stage. When the property was put up for sale by the South African Custodian of Enemy Property there was an estimated 434,000 tons of ore on surface dumps. The dump material, assaying 4.48 pct Cu, 15.8 pct Pb, and 10.9 pct Zn, had become highly oxidized during a stay of many years on the surface. But these dump ores had to be used in evaluating possible metallurgy because the shaft became flooded after the 1940 shutdown and underground ore was unavailable. Many companies examined Tsumeb but quickly lost enthusiasm because of the formidable metallurgical problems presented by the dump ore. However, on the basis of extensive test work, the present company decided that differential flotation would be successful and purchased the property after competitive bidding. Geology and Mineral Association is Complex The Tsumeb mine lies near the northern boundary of South West Africa at an altitude of 4200 ft. The village of Tsumeb is 335 air miles northeast of Wal-vis Bay, territorial port of entry. The orebody is a hydrothermal replacement deposit associated with an irregular intrusion of aplite cutting folded pre-Cambrian dolomite beds. In plan the deposit is a lens of 250x600 ft maximum dimensions. In section the orebody dips 55" south from the surface to 2000 ft, below which there is a reversal in dip to 75" north. Mineralization is localized in the aplite and in the adjoining brecciated dolomite. In the upper levels the ore is largely confined to a single lens, but below the 1890 level massive sulphide ore forms north and south lenses that partially enclose a pipelike body of medium to low grade disseminated ore, part of which is of too poor a grade to mine. Drilling below the 24th, or bottom level, has disclosed that a high percentage of the mineralization from the 3000 to 3300-ft horizons is comprised of copper and lead oxides. The deep oxidation may be local, since it appears due to circulating waters in a major fault cutting the orebody. Tsumeb ranks as one of the highest grade base metal deposits in the world. It is geologically and mineralogically famous for 41 copper, lead, zinc, vanadium, and germanium ore minerals, but its complex mineral assemblage is not attractive to metallurgists. In addition, there is a gradual change in oxidation intensity from the upper levels to the lowest level now operating, which is the 24th. This is well illustrated by the copper minerals. In the upper levels malachite and azurite are predominant, in the intermediate levels chalcocite, and in the lower levels tennantite. Throughout the mine there are sufficient heavy metal soluble salts to activate zinc, and large amounts of depressants are required even for so-called clean sulphide ores. When the same sulphide ore is exposed to prolonged oxidation aboveground, as was the case with the dump ores, reagent requirements become fantastic. Plant Construction First staff members arrived at Tsumeb in March 1947, and by May essential services had been restored and mine unwatering began. In 2 1/2 months company employees built a temporary jig and sorting plant for treating existing dumps to provide a saleable product during high metal prices. Meanwhile, concentrator design and construction was contracted to a South African firm, with subcon-

Jan 1, 1956

By Robert C. Emmett, Donald A. Dahlstrom

FOR many years gravitational classification has been employed as a basic tool in beneficiation of minerals and coal. While improvements have been made to increase efficiency and fields of application, the basic flow pattern has remained relatively unchanged in most gravitational classifiers. Feed slurry is introduced into a settling pool wherein the coarser solids settle against either an upward or horizontal fluid current, or a combination of both, and are withdrawn in a concentrated underflow stream. The finer solids below the classification point possess too low a settling velocity and thus are primarily swept out through a weir overflow by the bulk of the fluid. Because of its simplicity, efficiency, and low power and maintenance costs, gravitational classification finds a wide acceptance for separations in the range of 20 to 325 mesh. Two factors resulting from the basic flow pattern employed in gravitational classification present difficult problems for certain installations. If classification at 60 to 200 mesh of separation is desired, capacity in terms of gallons of overflow per minute per square foot of settling pool area must be low to allow required solids to settle to the underflow. If a thousand or more gal per min of feed slurry must be handled, either a very large classifier or a number of smaller units must be employed, taking up floor space and head room. Examples of this are the large classifiers, wetting tanks, or settling tanks used in the coal industry for recovery of process water from fine solids. Units as large as 50 or 60 ft in diam and 60 ft high, requiring substantial amounts of steel, are not uncommon. A second factor is the lack of self-regulation in gravitational classification. As overflow rate increases, mesh of separation increases correspond-ingly. This normally heightens difficulties of bene-ficiation. If a fine mesh of separation is required at varying feed rates, the problem is more than doubled in complexity. One of the primary advantages of gravitational classification is low power requirement. Gravity feeds or very low pressure heads are employed. However, fluid energy is still present in the inlet stream and provision must be made to dissipate it properly to obtain uniform upward or horizontal currents. Rather than waste this energy, it would be desirable to utilize it to secure a greater degree of self-regulation and simultaneously increase classification capacity per square foot of area. A free vortex could be realized by a tangential entry of feed, resulting in generation of centrifugal force to complement force of gravity in its action. With greater force it should be possible to gain finer separation at any overflow rate per square foot of settling pool area as compared with conventional gravitational classification. Furthermore, as feed rate increases, centrifugal force will also increase and at least partially offset faster currents sweeping solid particles to the overflow. A more stable classification point should result. The liquid-solid cyclone is the most common application of a free vortex in the beneficiation field.' Many previous papers have emphasized uniformity of flow pattern within the cyclone which is largely responsible for the sharp classification that can be obtained.'-4 If a free vortex can be maintained in a non-pressurized vessel with a free surface it may be possible to realize a third advantage of sharper classification. This open-top cyclone would then achieve a minimum energy loss comparable to that with conventional classifiers, and a mesh of separation coarser than 150 mesh, extending the range of application of the free vortex. This is not presently possible with ordinary cyclones. To test the applicability of the free vortex to gravitational classification, a 30-in. diameter open-top cyclone was constructed at Northwestern University. The cylindrical section was 30 in. high and the conical section included an angle of 60". The feed nozzle was a standard 4-in. pipe entering tan-gentially, with centerline 17 1/2 in. below the top of the cyclone. The conical section was flanged near

Jan 1, 1954

US 4,141,854-Method of resolving water-in-oil emulsions from the organic solvent extraction of uranium from an aqueous ore leach solution The emulsion is treated at optimum temperature with a water-soluble acrylamide copolymer which contains 5-50% by weight of a lower alkyl substituted tertiary aminoethyl methacrylate or quaternary ammonium salts thereof The emulsion is broken, enabling recovery and reuse of the organic portion of the emulsion A M Pavilcius and M A Latko, assigned to Nalco Chemical Co Feb. 27, 1979 3 pp (252-301 1W) Filed: 12-14-77 (860,542). US 4,142,952-In the recovery of copper from an ore leach solution, the solution is contacted with a water-immiscible organic solvent containing 5 to 20 g/L of o-hydroxyaryloxime and 2 to 20 times the weight of the o-hydroxyaryloxime of nonylphenol The loaded solvent phase is separated from the aqueous phase and stripped with sulfuric acid Copper is recovered rapidly and in high yield from aqueous solutions containing high copper concentrations R F Dalton, as¬signed to Imperial Chemical Industries Ltd Mar 6, 1979 6 pp (204-106) Filed 1-3-77 (756,067). Priority: United King¬dom, 1-30-76 (3788/76), 6-15-76 (24722/76), and 10-21-76 (43711/76). Same- British 1,549,615, dated Aug. 1, 1979 US 4,148,631-In the recovery of nickel and cobalt from a mixed sulfate ore leach solution, both metals are extracted by means of a cationic extractant, the loaded extractant treated with an aqueous chloride or thiocyanate solution to back extract the cobalt, and the cobalt then recovered by solvent extraction or ion exchange with an anionic extractant Recovery is effected in a relatively few extraction steps. F Babjak, V A Ettel, and K. R Dymock, assigned to Interna¬tional Nickel Co , Inc, Apr. 10, 1979 6 pp (75-101 BE) Filed¬7-27-78 (928,401). Priority- Canada, 12-28-77 (293,967) US 4,148,813-An improved extractant for use in the recov¬ery of nickel and cobalt from an ore leach solution comprises a water immiscible organic solvent system containing solubilized mercaptide ions, e g , dodecanethiol. Cobalt and nickel mercaptides solvated in the organic phase are easily stripped in the presence of carbon monoxide and an aqueous alkaline solution A. S. Rappas and J. P. Pemsler, assigned to Ken¬necott Copp

Jan 1, 1980

By Jerry H. Service, F. L. Partlo

The following investigation was undertaken to develop a method for determining with reasonable accuracy the depth of overburdens of 100 ft. or less. Seismic methods seemed to offer good possibilities. The problem differed from that of the familiar salt-dome location in that the depth of overburden to be determined is much less than is encountered in oil prospecting. This means that a much shorter time interval must be measured with accuracy. Refraction methods seem at present to offer greater possibilities than reflection. The air sound persists for from 0.1 to 0.5 sec., because of echoes, and seismic reflections coming during this time are difficult to detect. A seismometer sensitive to vertical vibrations only offers possibilities and has been used here with some success. It is hoped that the application of this type of seismometer to reflection shooting in this region may be investigated still further. The apparatus used consisted of a carbon-button seismometer followed by a three-stage, transformer-coupled, vacuum-tube amplifier and a General Radio string oscillograph. Time marks are provided by a 1000-cycle, 600-r.p.m., synchronous motor provided with five spokes which intercept the light to the camera at 1/50 sec. intervals. This motor is driven by a two-stage vacuum-tube amplifier of 3 watt output, actuated by a tuning fork. The power supply for the seismometer amplifier is one 2-volt storage and one 45-volt B battery. The motor amplifier is powered by a 6-volt storage battery and 6 to 225-volt, 15-watt motor generator. The temperature-frequency characteristics of the tuning fork have been determined and for ordinary working temperatures the error in time that would be introduced is less than M millisecond for time intervals up to 100 milliseconds. The impact is produced by a sledge blow for surface distances less than 100 ft. and dynamite for greater distances. Two steel balls with wires attached form an electric circuit through a 1-ft. piece of railroad rail upon which they rest. When the rail is struck by the sledge the balls rebound, opening the circuit and indicating the time of impact by a break in the oscillograph record line as at A, Fig. 3. When dynamite

Jan 1, 1934

By M. B. Ramachandra Rao

The spontaneous polarization surveys dealt with in this paper were carried out near Guddadarangavvanahalli (lat. 14° 17' N.; long. 76° 24' E.) in Chitaldrug district, which forms part of the Chitaldrug schist belt. The occurrence of gold, copper, lead and antimony ores has been noticed for several years but efforts to exploit them have been unsuccessful, therefore some spontaneous polarization surveys were made in the hope of finding indications of workable ore bodies. The country examined has a rugged topography, with numerous hills on which there is little vegetation. The climate is generally dry, the average annual rainfall of the tract being about 25 in., of which 17 in. occurs during the southwest monsoon. The geological formations of the area pertain to an ancient schistose series known as the Dharwar system (Archean). Locally, the rock formations have been divided into three series1—the chloritic series (or the Chitaldrug formation), the clay series (or the G.R. formation) and the trap series (or the Jogimardi trap). The chloritic series includes a complex of altered ancient traps and volcanic agglomerates intermixed with some schists of sedimentary origin, all of which are highly metamorphosed. The lithological types of this series are mostly distinguishable as greenstone schists, calc-chloritic schists, slaty schists, ochery and calcareous mica-chlorite schists, with subordinate bands of secondary limestone. The clay series (G.R. formation) is made up of shaly and argillistic schists, litho-margic clays, sandstones and brecciated ferruginous quartzites. The Jogimardi trap is mostly of the nature of an epidiorite (hornblende diabase), which is distinctly intrusive in the form of large laccolithic and narrower dikelike masses. In addition to this trap, there are other minor intrusives such as the chalybitic trap, quartz porphyry and quartz reefs. The topographical and geological features of the area are noted on the map of Fig. I. The locations of some of the spontaneous polarization centers determined by the geophysical survey are indicated also. Method of Survey and Apparatus Employed The spontaneous polarization surveys were carried out in the manner recommended by the Imperial Geophysical Experimental Survey.2 Preliminary reconnaissance was first carried out along a series of parallel traverse lines that were laid out over the area. The Potential profiles obtained in these traverses were plotted* on a suitable scale and when any

Jan 1, 1946

By M. B. Ramachandra Rao

The spontaneous polarization surveys dealt with in this paper were carried out near Guddadarangavvanahalli (lat. 14° 17' N.; long. 76° 24' E.) in Chitaldrug district, which forms part of the Chitaldrug schist belt. The occurrence of gold, copper, lead and antimony ores has been noticed for several years but efforts to exploit them have been unsuccessful, therefore some spontaneous polarization surveys were made in the hope of finding indications of workable ore bodies. The country examined has a rugged topography, with numerous hills on which there is little vegetation. The climate is generally dry, the average annual rainfall of the tract being about 25 in., of which 17 in. occurs during the southwest monsoon. The geological formations of the area pertain to an ancient schistose series known as the Dharwar system (Archean). Locally, the rock formations have been divided into three series1—the chloritic series (or the Chitaldrug formation), the clay series (or the G.R. formation) and the trap series (or the Jogimardi trap). The chloritic series includes a complex of altered ancient traps and volcanic agglomerates intermixed with some schists of sedimentary origin, all of which are highly metamorphosed. The lithological types of this series are mostly distinguishable as greenstone schists, calc-chloritic schists, slaty schists, ochery and calcareous mica-chlorite schists, with subordinate bands of secondary limestone. The clay series (G.R. formation) is made up of shaly and argillistic schists, litho-margic clays, sandstones and brecciated ferruginous quartzites. The Jogimardi trap is mostly of the nature of an epidiorite (hornblende diabase), which is distinctly intrusive in the form of large laccolithic and narrower dikelike masses. In addition to this trap, there are other minor intrusives such as the chalybitic trap, quartz porphyry and quartz reefs. The topographical and geological features of the area are noted on the map of Fig. I. The locations of some of the spontaneous polarization centers determined by the geophysical survey are indicated also. Method of Survey and Apparatus Employed The spontaneous polarization surveys were carried out in the manner recommended by the Imperial Geophysical Experimental Survey.2 Preliminary reconnaissance was first carried out along a series of parallel traverse lines that were laid out over the area. The Potential profiles obtained in these traverses were plotted* on a suitable scale and when any

Jan 1, 1946

By W. P. Lass

The purpose of this article is not only to describe the plant and method of cyaniding the Treadwell concentrates, but to present some of the results of the experimental work obtained in the past three years for the Alaska-Treadwell Qold Mining Co., at Douglas Island, Alaska, under the direction of F. W. Bradley, Consulting Engineer, and Robert A. Kinzie, General Superintendent, of the affiliated companies. At the time the experimental work was undertaken the concentrates were being shipped to the smeltery at Tacoma, Wash., and the cost for treatment of 3-oz. (gold) concentrates was $11.95 per ton, divided as follows : Smelting-charges,......... $4.00 Loading, freight, insurance, etc., . 2.89 Interest due to time loat in transit and in settlement,... 0.05 Loss due to settlement for 95 per cent. of the gold at $20 per ounce, 6.01 Total,...$11.95 From the experimental work described later, it was estimated that 96 per cent. extraction could be made by treatment on the ground, and that the cost, when treating 80 tons per day, would be $3.25 per ton, divided as follows : Per Day. Per Ton. Labor,. $86.16 $@.827 Chemicals,.76.60 0.960 Power and steam-heat,... 67.60 0.845 Marketing-, refining- and other charges, . 49.36 0.617 Totals,.. $2.59.72 $3.250 Adding to this total the 4 per cent. treatment-loss, which on 3-02. concentrates amounts to $2.48, gives a total cost of $5.13 per ton. Comparing this with $11.95, the cost when shipping to the smelter, leaves a net gain of $6.22 per ton by the local

Jan 1, 1912

By N. C. Fick, R. L. Rickett

DESPITE the large volume of literature on alloys of iron and silicon,1 there is little published information dealing specifically with the constitution, at various temperatures, of the alloys containing less than 0.10 per cent carbon and from 0.5 to 4.5 per cent silicon that are used extensively in the electrical industry. Because such information was needed by those engaged in the production of these alloys in, sheet and strip form, the constitution was determined of a series of commercial silicon steels containing 0.05 to o.08 per cent carbon and 0.20 to 4.31 per cent silicon, over the temperature range 1300° to 2500°F. These alloys were all from regular production heats and of commercial quality. The results of, the tests are given in Fig. I and for comparison the constitution diagram of iron-silicon alloys with 0.01 to 0.02 per cent carbon, taken from the literature, is also presented. MATERIAL AND PROCEDURE All steels were hot-rolled to a thickness of 0.060 to 0.080 in. and used in this condition. Their chemical compositions, including over-all check analysis, and "center" analysis for carbon and silicon, are given in Table I. The so-called center analysis was determined on samples pickled to remove the outer surface and leave only the center third of the gauge thickness-the portion of the sample on which metallographic ratings were based. In most, but not all, of the alloys, slightly higher carbon was found at the center than when the whole strip was sampled; the two methods of sampling show relatively little difference in silicon. Two of the steels, A and E, contain relatively high phosphorus, while I and K are somewhat higher in manganese than the others. [ ] The procedure consisted essentially of heating specimens (about 1/16 by 3/8 by 3/4 in.), protected from oxidation or change in carbon content, at a temperature in the desired. range and for a selected period, after which they were quenched in iced brine; the degree of transformation that had then occurred was determined by metallographic examination. The soaking periods were selected after preliminary investigation had shown that a con-

Jan 1, 1946