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By Vincent P. Gi. ccnella

GOLD was discovered in Gold Canyon on May 15, 1849. Following this discovery placer miners worked the gravels in the canyon for-ten years, finally discovering the outcrop of the Comstock lode at Gold Hill in May, 1859, and in Ophir ravine on June 12 of the same year. Before long large quantities of gold and silver, were taken from the lode and one of the West's most spectacular mining booms drew thousands of fortune seekers to Nevada. After a long period of production activity waned and in late years little mining was done. The recent reopening of the old mines and the discovery of ore in other parts of the vein, and in other veins of the district, has led to a revival of interest in the Comstock. Virginia City, Gold Hill, and Silver City, situated along the great vein system, are enjoying a period of prosperity and a large influx of people.

Jan 1, 1934

By George A. Moore

A recent survey of existing analytical results, and an attempt to correlate them with each other and with the known history and behavior of the samples, indicated that none of the commonly applied methods could be considered established as intrinsically reliable. It was. therefore, desired to attempt to establish the reliahility of some one method by external evidence. If this were to be accomplished, other methods could then be evaluated in terms of the method establibhed. The correct method was accordingly to be chosen without regard for cost, speed, or convenience; with the requirements for routine application to be reserved for 1ater consideration. The method of total combustion was chosen and submitted to a limited number of tests against primary standard samples. Considerable experimental difficulty was encountered, indicating that additional work on the experimental arrangements is desirable. While the work to (late by no means represents a completed investigation, it is helieved that the combustion method is basically sound and that its development should be continued. Choice of Method Of the several available methods, mechanical disintegration. electrochemical methods'2'3 modifications of the Sieverts experiment,% and vacuum fusion, were all judged to contain one or more factors of un- certainty not at present completely cvaluated or suhject to complete control. These methods, therefore, were considered currently unsuitable for referee purposes. Solid evolution was considered to be of doubtful significance when conducted in the alpha range. A reasonable chance for a true analysis was found for evolution in the gamma range. Preliminary tests have indicated a very slow rate for evolutioll from gamma iron. indicating a running time of possibly 100 hr. for a one-centimeter sample. Difliculty in distinguishing such slow. evolution from blank gas was anticipated, hence the method was not consitlerctl the most suitable for the present purpose. The combustion method was evaluated as being most sound on fundamental principles. The only important unknown would be to determine that the iron oxide formed did not occlude water at the tcmperaturv of the combustion.with the odvious precautions of using a sound and clean specimen, providing a system sufficiently tight to give acceptable blanks in longtime runs, and providing a suficicntly sensitive measuring system, reasonable, precise, and truthful analytical results could be logically expected. Apparatus The equipment as originally operated is shown in Fig. I, and the glassw-are in its latest modification in Fig. 2. The sulphuric acid drier and small safety tube in the upper right of Fig. I were replaced by a KOH drier and combined manometer safety tube, to permit operation at reduced

Jan 1, 1945

By George Otis Smith

A YEAR ago, ,at the Institute's dinner, I closed my A remarks with the words: "The scientist devotes his life to the advancement of learning; the engineer gives his to the advancement of living." May 1 begin tonight where I left off then and speak first of the distinction between science and engineering? This difference is a difference both in purpose and in method: it is the sharp contrast between theory and practice; it marks the broad gap between dream and attainment. Sometimes it happens that the investigator is also the builder and applies what he discovers, yet the distinction between science and engineering serves to point out the broad dependence of the engineer upon his brother the scientist. You remember .the old - saying that "where there is no vision the people perish" translate this into modern experience and we may say that "where there is no science, the engineer fails."

Jan 1, 1929

By Reno Sales

INTRODUCTION. THE geology of Butte possesses especial interest on account of the magnitude of the ore deposits, their extraordinary richness and persistence in depth. Since its discovery in the early 60's the district .has yielded, in round numbers, 6,000,000,000 lb. of copper, 260,000,000 oz. of silver, 1,250,000 oz. of gold, and a large but not definitely known number of pounds of zinc. This enormous metal production has been derived from an ore tonnage estimated roughly at 65,000,000 tons. At the present day, most of the important mine shafts have reached a depth of. 2,400 ft., while several are more than 3,000 ft., deep. At these-great depths rich bodies of copper ore are being exploited, showing no diminution in metal content as compared with the ores of the higher levels. The present (1913) daily rate of production is approximately 14,000 tons, of which 13,250 tons may be regarded as copper ore, although a small percentage of this is minable only because of the presence of

Jan 8, 1913

By John J. Collins

[CONTENTS PAGE Purpose and scopeI Core samplingI Sludge sampling.....7 Combining core and sludge assays13 Deviation and surveying of drill holes17 Conclusions24 Acknowledgments25 Bibliography25] PURPOSE AND SCOPE The purpose of this paper is to review a few of the problems involved in the interpretation of diamond-drill core and sludge samples, with the hope that it will stimulate discussion and bring forth some of the information on the subject that has been accumulated by various organizations. The release of these data would benefit the mining industry as a whole. Particularly valuable would be the details of drilling programs that have been tested by actual mining operations, as most of the available published material is unsupported by production records. The scope of this paper is limited to a few of the commonest problems of interpreting diamond-drill core and sludge samples and the deviation of holes. It does not include the description of routine operations performed by the drillers, nor the estimation of ore reserves. The methods of calculating reserves have been described 25,26,31,32.50,65 but more information is needed on the results shown by actual mining operations. CORE SAMPLING The accuracy of diamond-drill-hole sampling depends in large measure on the percentage of core recovery. As a general rule, the longer the run, the poorer is the recovery. Although the longer runs allow cheaper drilling costs, the determining factor should be the percentage of recovery desired. The problem is one of balancing accuracy against costs. Holes of larger diameter may show improved recovery, except as Joralemon* points out that there are some conditions in which small holes give better core recovery than large holes. This is particularly true, he says, in rocks like shattered quartzite, where larger core tends to break in angular fragments and grind up, while the diameter of smaller core is less than the distance between joint planes and grinding is not so serious. In a large measure the length of the sample run should depend on the type of deposit. Favorable drilling may allow uniform sample runs of 5 ft., 10 ft., or perhaps 20 ft. In veins or beds of relatively narrow widths and changing characteristics, however, the ore zone may be overrun and ground in a core barrel full of waste rock. Generally it is desirable to break the sample at a change in the rock or ore. Such a change, often indicated by a color change in the sludge water, should be a signal for the driller to make a "dry" block and to pull out that run regardless of the length. Such a procedure permits the segregation of waste from ore samples so

Jan 1, 1946

By B. F. Tillson, C. M. Haight

I. General Remarks..........................723 1. Location............................723 2. Characteristics of the Orebody..................725 (a) Mineralogical (b) Shape, Strike, Dip, Size, Etc. (1) Ore (2) Country Rock 3. Historical Sketch........................725 11. Mining Method..........................726 1. Outline of Methods Now in Operation............... 726 (a) Stopes (b) Fill . (c) Pillars (d) Slices (e) Transportation (f) Miscellaneous 111. Mining Operations........................727 1. Shaft..............................727 (a) Sinking (b) Dimensions and Equipment, Sketches (1) Size, Inclination. Direction, Etc. (2) Timbering (3) Stations (4) Track (5) Hoisting Equipment (a) Skips (b) Engines and Ropes (c) Idlers (d) Man Cars (6) Loading Pockets (a) Operating Figures (7) Headframe

Jan 1, 1918

By George Harrington

THIS paper is a brief description of the design and equipment of a new coal mine in southern Illinois, which has many features not common practice in shaft coal mining and which is laid out and equipped for such unusually large capacity that it promises soon to become one of the largest mining operations of its kind. Shaft sinking for the mine, New Orient, was commenced in May, 1921, by the Chicago, Wilmington & Franklin Coal Co. In October, 1924, the plant was practically completed and was producing over 6000 tons of coal per day. There is good prospect that within a few years, when the mine is fully developed and manned, the output will attain a daily rate of 12,000 tons. The coal seam at New Orient lies approximately horizontal, about 500 ft. below the surface, and may be said to run from 9 to 12 ft. in thickness. The tendency in Franklin County has been toward large mining units (5000 to 6000 tons per day) with great care being given to protection against fire and explosion hazards, and usually with quite elaborate surface plants permitting careful preparation of the coal into graded sizes comparable with those made in the anthracite region. The field is unionized, with uniform conditions of wages for various classes of work, hours, etc., and the Illinois mining laws are quite specific with regard to fireproofing and safety measures. These conditions make it difficult for any but well-equipped and well-operated mines to meet the competition of the numerous fine properties of which this county can justly be proud. The coal company's property was already developed on the western side by the Orient. mine-a much visited and written about operation (with which many of you are possibly familiar through the Department of Mines' film "The Story of Coal ") which until recently held the 8-hr. production record with 8218 tons and usually ran quite consistently with production of slightly over 6000 tons per day. Experience at Orient was naturally one of the greatest influences in determining the location and character of the New Orient.

Jan 2, 1925

By Roy E. Dickerson

CUBA differs considerably from the other Greater Antilles in many geologic fundamentals. Cuba is geosynclinals; whereas Jamaica, Hispaniola. (Haiti), and Puerto Rico are geoanticlinal. (Scliuchert, Chas., Historical Geology of the Antillian. Caribbean Region, p. 13, 1935.) Rocks of lowermost Cretaceous age exist in large volume in Cuba, where- as they are generally absent in the other islands. Since many of the surface indications of petroleum are intimately associated with lowermost Cretaceous rocks, the Vinales limestone,, this difference is most important. The Vinales limestone in many places in Pillar del Rio and Santa Clara provinces reeks with light tars when broken with a hammer. The late David White and his associates on the U. S: Geological Survey studied

Jan 1, 1937

By Archer E., Wheeler

COPPER operations in the Congo mean the operations of the Union Miniere du Haut Katanga, because there are no other copper industries there. There is a mine at Bwana M'Kubwa, a little way to the South, in Rhodesia, which has been operated on a small scale, and-another farther south, the Falcon mine, but in quite a different section of the country. This Bwana M'Kubwa mine has recently been reorganized financially and the Minerals Separation interests have taken an extensive share in it, so that it may come into operation. The Union Miniere's operations are in Katanga, the most southeastern province of the Belgian Congo, and the company's concession is in the southeastern part of that province. The usual route for going into the country is by way of Cape Town, thence over the South African Railways through Kimberly and Mafe- king to Bulawayo, then by the Beira, Mashonland and Rhodesia Railways to the Rhodesia-Congo border, passing Victoria Falls on the way, and finally by the Belgian road, the Chemin de Fer du Katanga, to Elisa- bethville, the African headquarters of the company. From Cape Town to the border it is about 2150 mi. and from the border to Elisabethville about 158 mi., making a little over 2300 mi. from Cape Town, by rail.

Jan 1, 1924

By Claiborne C. Van Zandt

THE heavy postwar demand for Portland Cement has created shortages that are gradually being overcome by increases in plant capacity. In the post-war period, the Lone Star Cement Corp. has expanded several of its existing 15 plants and has built 2 new plants, now in operation, with a third new plant under construction in Brazil, of design somewhat different from that of the two preceding ones. The new domestic plants are located at Lone Star, Va., near Roanoke, and at Maryneal, Tex., near Sweetwater, and went into operation during 1951. Each plant is equipped with two rotary kilns and has an initial capacity of 1,500,000 bbl per year, or 6,000,000 sacks. The total cost of each plant was about $10,500,000, and on this basis $7.00 in 1951 money are required per bbl of annual capacity; at 376 lb per bbl, this corresponds to $37.00 per ton of output. The investment corresponds to approximately $80,000 per hourly-wage earner employed. The purpose of this paper is to describe the most interesting design features of the two new and almost identical Lone Star plants. Why the Dry Process Was Chosen The two new domestic Lone Star mills employ the dry process because: 1-The raw materials at both properties are particularly suitable and can be dried very, readily. 2-The Maryneal, Tex., plant does not have water available for slurry. 3-Very substantial kiln fuel savings can be obtained. 4-Dry raw grinding reduces ball and liner wear and- general maintenance, without any sacrifice in power. Operating costs to date have more than justified the choice of the dry process for these two plants. Raw Materials The chemical composition of Portland Cement raw material after crushing, grinding, and mixing, but before burning, is about as follows: [ ] sometimes tailings from an iron ore concentrating plant. Certain objectionable materials occur in raw mix: magnesia, limited in cement to 5 pct, in excessive amounts; alkalies, as sodium and potassium; phosphorous; and sulphur, generally present in coal and oil. The Roanoke raw, materials are dense limestone and shale. The Maryneal materials are soft limestone and a hard brittle clay that can be handled much like shale. Sand and iron ore are required in small quantities. General Plant Layout and Construction The two plants are so similar that one joint description will cover both, with such exceptions as will be especially indicated. The crushing plant is located in the quarry, and raw material is conveyed by belt to the storage building, where all materials are stored and handled. Fig. 1 identifies the various structures of the Roanoke plant. The large storage building is used for the storing and handling of all materials, including limestone, shale, sand, iron ore, coal, clinker, and gypsum. Directly adjoining the storage building is a structure under a similar arched roof, containing the raw and clinker grinding departments and the electrical substation. The discharge ends of kilns and coolers, and coal or gas burning equipment are also placed at one end of the mill building. The 340ft long kilns extend away from the main buildings to form the long leg of an L, at the end of which is the feeding and kiln dust collecting equipment, and the raw blending silos are conveniently located adjacent to the feed end of the kilns. The machine shop, storeroom, electrical shop, foremen's office, service rooms, and meeting hall are combined in one structure in the center of the plant, and the office and laboratory are close by. The cement silos and packhouse do not have to be tied directly into the manufacturing process and are therefore located at a point convenient to the railroad tracks.

Jan 1, 1952

By R. J. Teitel

FOUR or five years ago an investigation was initiated to study and develop basic information on llquid metal fuels for future power nuclear reactors. The objective was to find fluid forms of uranium (solutions or dispersions) in low melting alloys. A summary of the different fuels was presented in a recent article by Teitel, Gurinsky, and Bryner.' The combination of uranium and bismuth, with its low thermal neutron cross section and low melting point, is of basic interest to the development of these fuels. Most of the previous effort on the U-Bi phase diagram was done under security regulation. Ahmann and Baldwin' reported their findings on the U-Bi system in 1945. Their results indicated that the phase diagram had at least two compounds and possibly one more. UBi, and UBi were established by chemical analyses of residues after high temperature centrifugation of alloys. UBi, decomposed peritectically at1980°C and there were signs of another reaction above that. UBi melted at a very high temperature, and there were also signs of a monotectic at high temperatures in high uranium alloys. The solubility of uranium in liquid bismuth has received attention by three groups of authors: Bareis,X ayes and Gordon,' and Ahmann and Baldwin.' The latest and most widely accepted data is that of Bareis, and it is included in Fig. 1 along with new data at higher temperatures. Brewer et al." and Ferro"'' have used X-ray diffraction to determine the number and crystal structure of the compounds. Brewer identified the structure of UBi to be a simple NaCl cubic structure having a cell edge of 6.364A. The density of the compound was calculated to be 11.52 g per cu cm. Two other unindexed patterns were presented as possibly belonging to two other compounds. Ferro agreed with Brewer on the crystal structure of UBi and identified the structure of UBi, and U,Bi,. The intermetallic compound UBi, has a tetragonal cell with a = 4.445A and c = 8.908A. The density of the compound was determined and the number of atoms per cell (six atoms) was calculated. The X-ray density reported was 12.35 g per cu cm. The compound UsBir was treated in a similar way and it was reported to have a cubic structure 9.35A on an edge and to contain 28 atoms per cell. The density was given as 12.59 g per cu cm. Teitel reported that there was a compound UaBi, and no UBi, compound. (The difference in the compositions is 3.3 pct U.) This conclusion was based upon the extrapolation of some 20 tie lines in the

Jan 1, 1958

By R. F. Yap, N. Arbiter, C. C. Harris

A fully-instrumented driving mechanism has been constructed to study the power, aerating and solid suspension characteristics of several laboratory flotation machines. Machines operating over normal flotation speed ranges give substantially constant power numbers in liquid systems indicating that they operate close to fully baffled turbulent flow conditions. Owing to lack of geometrical scaling, power numbers for different sizes of cells of the same make are different. Even larger differences occur between cells of different manufacturers. For a given impeller, varying the tank size did not significantly affect the power consumption. Tank geometry and baffling has a slight effect on power consumption with shrouded impellers. For a given impeller speed, suspension height decreases while power increases with increasing mass of solids in suspension. While power consumption is hardly affected by particle size and clearance between the impeller and tank bottom, the suspension height significantly increases as clearance increases and as particle size decreases. Aeration was found to affect not only power but also the suspension characteristics of the machine. For closely sized particles, drastic sedimentation occurred when a critical value of the air flow number was exceeded. The critical air flow number decreased with increasing particle size. This effect was less marked with a wide particle size range. The self-aerating capacity of flotation machines increases with cell size at a given speed for the same impeller-shroud system, but decreases in the presence of frother and solids. Equations have been established for correlating all of the studied variables. Data taken from the literature were also correlated with these equations. The modern mechanical flotation machine has, over a period of 60 years, displaced other designs to become the dominant type. Once the correct chemical environmental conditions have been established, the major requirement, selective mineral-air adherence, is essentially a function of two operating variables — impeller speed and air rate. The special nature of this dependence is such that in the same apparatus, but under different operating conditions, adherence can be either prevented or destroyed. Another aspect of this observation is that flotation machines though differing in detail to a considerable extent can produce similar metallurgical results. It must be concluded that having determined the correct reagent levels, establishing the correct operating conditions are at least as important as machine design. Since about 1960, after 30 years of fairly stable circumstances, manufacturers have been developing new types of machines whose major feature is emphasis on pulp flow characteristics by virtue of new im-peller-stator designs. As far as can be ascertained, these new designs necessitate radical departures from accepted impeller speed-air rate values. Performance data are still awaited. Flotation machine development has been empirical, as is selection of operating conditions. The purpose of the work at Columbia University has been to elucidate the fundamentals of the subject so that the purely mechanical aspects of the flotation process can be understood in hydrodynamic terms.1-6 Hopefully, this might lead to intelligent selection of a particular machine and its operating conditions for a specific ore based on a minimum of testing, or even to new types of flotation machines. The present investigation is the most comprehensive to date.7,8 Earlier work, which was largely exploratory, has been repeated and extended and employes the most accurate measuring techniques available. Additionally, sufficient hydrodynamic data were obtained under flotation conditions to delineate the relevant operating ranges. Attention has been given to scaling in that several geometrically scaled tanks and impellers have been constructed and tested; it has been pointed out elsewhere that undistorted geometrical scaling rarely pertains in flotation machine design.5 This paper is divided into sections dealing with relevant combinations of the three phases: liquid, air and solid. Summaries of the literature have been given in previous publications2-4 and will not be repeated here. Recently, several reviews of mixing have appeared, which can be usefully referenced.9-11 It is of interest to note that flotation is not included

Jan 1, 1970

By L. F. Pett

ALTHOUGH Kennecott's orebody has long been outlined, it is still necessary to define further its limits. This mine, long an advocate of churn drill methods, recently supplemented its practice by using a diamond drill. So far, the diamond drill holes have been along the main canyon floor which traverses the mine area. This zone, highly fractured, is cut by numerous friable quartz veinlets encountered within the porphyry. These veinlets, formed by filling contracted porphyry fractures with hot, siliceous, mineral-bearing solutions, are often more highly mineralized than adjacent porphyry. They pulverize on contact, yielding only slight amounts of core.

Jan 7, 1951

By John M. Campbell, W. E. Portman

A series of correlating charts have been prepared to enable the field engineer to predict the amount of stock tank fluid produced by stabilization of first stage separator fluid. The charts shown are particularly convenient for estimating the relative effect of temperature and pressure in economic studies. Further work is planned to investigate the factors affecting stage separation' yo supplement this preliminary work. The direct correlation of separation variables is dictated as a convenience to eliminate the trial and error calculation necessary when using the standard flash vaporization equations. These are obtained by a material balance using the equilibrium vaporization ratio K, and have appeared in the following, or similar. form many times in the literature. where La = mols of any component n in the separator liquid Ko — equilibrium vaporization ratio for any component at separator pressure and temperature Fa = mols of any component n in the feed Vn = mols of any component n in the vapor V = total mols of separator vapor These equations may of course be effectively and accurately solved by use of digital and analog computers. However, these are not usually readily available in field offices and the need often arises for fast results during routine operational problems. Consequently, many methods have been devised in an attempt to simplify the basic calculation.2-14 The methods proposed are generally satisfactory, but do not eliminate trial and error methods and the necessity of subsequent calculations to obtain the answer in volume terms. The purpose of this work. therefore, was to develop accurate charts that would give the result directly in gallons. When using stage separation the recoveries would be lower than those shown, for stabilization usually gives about 95 + per cent recovery on the potential stock tank components in the separator liquid. The recoveries are shown in terms of a 14 psi Reid Vapor Pressure product which has a true vapor pressure of 14.7 psia at 98°F. Consequently, this product represents a stable stock tank liquid under normal storage conditions. To estimate natural gasoline re- coveries by stabilization the volume shown may be multiplied by 1.20 for 18 Ib RVP natural gasoline and 1.36 for 26 lb RVP natural gasoline. PROCEDURE The five analyses shown in Table 1 were chosen since they were representative of wellstream compositions often found and had almost the same propanes plus liquid content, expressed in gallons per 1,000 std. cu ft of wellstream. The distribution of the components though is widely variant. A temperature range of 0-100°F. and a pressure range of 200 to 1,250 psia was chosen inasmuch as most separations fall within these limits The K values used were modified values of those originally published by Katz and Hachmuth.' In the range chosen these have generally given good results when predicting fluid recoveries. The choice was somewhat arbitrary, for the authors' personal experience has shown that

Jan 1, 1957

By Dudley L. Davis, E. C. Peterson

INTRODUCTION THE Darwin District is 30 miles east of Olancha which is 220 miles north from Los Angeles via U. S. Highway No. 6. The ore deposits occur in the Darwin hills that have been elevated above the Darwin Plateau by typical basin and range-type faults. Some of these faults have been active quite recently as shown by the displacement of Quaternary (?) basalt flows and the rejuvenation of erosion in Darwin wash. The ore bodies occur as fissure fillings and replacement of favorable beddings in a thick series of Paleozoic limestones, dolomites, shales and quartzites. Rich oxidized lead-silver ores were discovered in the early seventies. By 1875, Darwin had a population of about 5000, and by 1880 several small mills and smelters had been built. Early exhaustion of surface ores, the isolated location of the district, and fluctuation of metal prices caused intermittent operations until World War II. The district production is estimated at about $3,000,000 prior to 1900 and perhaps $4,000,000 between 1900 and 1945 when Anaconda purchased the major producing mines. For the past two years, daily production has averaged 75 tons of direct shipping and 150 tons of milling-grade ore. At present, 300 tons of mixed oxide and sulphide ore are treated by flotation, the oxidized lead minerals being activated by addition of sodium sulphide after galena and sphalerite have been recovered from the circuit. Extraction averages 85 pct of the lead, 80 pct of the silver, and about 33 pct of the zinc in a bulk concentrate. The mill is designed so that minor changes permit selective flotation of pyritic lead-zinc ores encountered at depth. GENERAL GEOLOGY The oldest rocks in the Darwin hills are a series of Paleozoic limestones, dolomites, shales and quartzites probably Pennsylvanian in age. These rocks have been intruded on the west by the Coso granite batholith and on the east by a granodiorite stock. Numerous sills and dikes grading from orthoclase granite to gabbro are exposed in the underground workings. The sedimentary rocks have been considerably folded and faulted, the most prominent structure being a northwest pitching anticlinal fold, the crest of which lies just west of the ridge line of the Darwin hills. The east limb of this fold has been intruded by the granodiorite stock and east of the stock the sediments show several closely spaced anticlines and synclines. The minor flexures on the west flank of the major fold are structurally important in localizing the ore in the Defiance and Essex mines of the Darwin group. There are three principal systems of faulting and fissuring, most of which are post-intrusive and premineral in age, with minor amounts of postmineral movement. The Darwin tear fault is the largest, most persistent fault in the district. It strikes N 60 to 95°W, dips steeply south and is traceable for about 1o miles. It is a normal

Jan 1, 1947

By H. B. Pulsifer

THIS paper outlines a method for rapid production of flat, granular surfaces on many of the medium hard- alloys. Grinding wheels and fabrics on wheels are not used; the purpose is accomplished by rubbing on emery papers and on fine abrasives heaped on hand boards. The advantages of chromic acid in the solutions for etching nickel and copper alloys are explained. Steels may require an additional step in the mechanical smoothing owing to the lack of an ideal etchant. Chloric acid is a. rapid etchant for steel. Thirty-six photomicrographs illustrate the steps in smoothing and the clearness with which structures can be rapidly exposed. New structures in low-carbon steels are announced. After a little practice, many materials can be finished within two or three minutes. CUPRONICKELS In the previous papers on the metallography of copper and magnesium which the author has presented to this Institute there has been an attempt to demonstrate the advantages of minimizing the mechanical work of surfacing and depending more on chemical means; the present paper continues the research into a still wider field. It is obvious that the success of a chemical method for smoothing and etching depends almost entirely on the use of a suitable reagent. The etchant must remove the surface granulations and outline the structures. A very special type of reaction between reagent and metal must be displayed. Promiscuous corrosion or the intensification of the attack about local couples -formed by impurities or components must not be in evidence. There are several reagents that etch highly polished nickel and copper alloys very well. Acidified ferric chloride solutions, concentrated and dilute nitric acid solutions, and nitric and acetic acids in acetone have -been used to good advantage for the purpose. However, if they are expected to etch rather deeply-that is, smooth as well as characterize-the results may be far from satisfactory.

Jan 1, 1928

By J. A. Morgan, R. E. Lund, D. E. Warnes

The design, operation, and performance of an integrated pilot plant for recovering zinc and copper from a complex sulfide ore are described. Metallurqical processing comprised selective sulfate roasting in a fluid bed, leaching in a weak sulfuric acid solution, recovering copper by cementation with scrap iron, air oxidation to remove soluble iron, more-or-less conventional Purification for removal of impurities deleterious to zinc electrolysis, recovering zinc by electrolyzing solutions, and recycling spent electrolyte to the sulfating roaster. The effects of several feed preparation and roasting variables on the sulfation response of zinc, copper, and iron are set forth, together with the electrolyzing characteristics of the zinc solutions. SEVERAL years ago, St. Joseph Lead Co. undertook a research program to develop procedures for winning values from a complex ore. The deposit was a massive sulfide ore body containing upwards of 65 pct pyrite. Gangue, composed principally of quartz, calcite, dolomite, and alumina, comprised only 10 to 20 pct of the mineralized zones. Zinc, lead, and copper values were variable, but averaged about 6.7 pct Zn, 2.6 pct Pb, and 0.4 pct Cu. Early in the exploratory test work, it was established that the intimate physical association of the minerals in the ore made beneficiation by conventional flotation procedures difficult. Accordingly, in addition to a comprehensive program to investigate physical methods of ore beneficiation, a research program was established to develop a direct metallurgical method for processing the ore. After surveying a number of potential methods for recovery of both metal and sulfur values, effort was concentrated on the development of a sulfate roasting process. It is the pilot-plant development of the sulfate roasting-solution purification—electrolyzing procedures which forms the subject of the present paper. PREPILOT PLANT INVESTIGATIONS From a thermodynamic viewpoint, it is feasible to sulfate selectively a host of metals, including Zn, Cu, Pb, Cd, Ni, and Co, in an iron sulfide ore without sulfating iron. Of course, a number of conditions such as roasting temperature and gas composition must be controlled. In addition, as first disclosed by Hybinette,' an alkali metal salt is frequently beneficial in promoting the sulfation reactions. The application of fluid bed roasting to the sulfate roasting of nonferrous metals was pioneered by Stephens of Battelle Memorial Institute.' Subsequently, Falconbridge Nickel Mines, Ltd. developed an unique process for selectively sulfating and recovering nickel, copper, and cobalt from their nickeliferous pyrrhotite, which recently has been described in an excellent paper by Thornhill.3 One of the unique features of the Falconbridge process is that the feed is introduced into the fluidized roasting furnace in a manner such that the feed slurry is broken up into droplets and the droplets lose their moisture before contacting the surface of the fluidized bed. The dried droplets do not tend to cake or fuse, but rather remain as discrete agglomerates of concentrate particles and roast without significant degradation. St. Joseph, aware that Battelle Memorial Institute had participated in the development of the Falconbridge process, obtained permission from Falcon-

Jan 1, 1962

By J. L. Perry

FOR many years, you and your fellow members of the American Institute of Mining and Metallurgical Engineers have devotedly and ably applied yourselves to the art of making iron and steel. having foremost in your minds the conservation of raw materials, suitable and efficient facilities, economical processes, and improved quality of product. For the past three strenuous years you have contributed wholeheartedly to meeting the problems of increased production. You have conducted scientific and technical investigations and research, recommended experiments. and gathered all the available and valuable information possible to gain the desired objectives. You have had many obstacles and difficulties to overcome. In the early days of your organization, there was some lack of co-operation from the industry itself-perhaps, for example, some failure to recognize the mutual advantages of free exchange of experiences and information. But regardless of any hindrances, you have made many and valuable contribution which have tended to make iron and steel manufacture leas of an art and more of a science.

Jan 1, 1944

By EDWARHD HALSE

Discussion of the Paper of Dr. H. M. Chance, presented at the Lake Superior Meeting, September, 1904. EDWARD HALSE, Puerto Berrio, Colombia, So. Amer. (communication to the Secretary*) : The Taviche mining-district is generally known as a Taviches," being named from the two villages called respectively San Pedro Taviche and San Jeronimo Taviche. It is so called by José G. Aguilera, in his paper on the Geographical and Geological Distribution of the Mineral Deposits of Mexico;' and in published reports by the late N. W. Rudston Read ;2 and by Juan de Dios Villarello.3 The region is a large one, measuring about 24 miles long and 7 miles wide. Dr. Chance, in his interesting paper, tells us very little about the nature of the ores, and, with regard to the geology of the district, contents himself with observing that ? the country, as a rule, consists of eruptive rocks; and no evidences of sedimentary strata appear near the mines.? The latter part of the above statement is hardly correct. The formation is mainly eruptive, being hornblende-audesite of Tertiary age, which is limited on the east by the calcareous (Cretaceous) formation, and on the west by the Huronian rocks of Ocotlan. Limestone is seen close to the Santa Càtarina mine, at the northern end of the region. In the surface-working known as La Cueva, a little northeast of San Pedro, limestone forms the eastern wall of the deposit which consists of spots (moscas) of pyrargyrite in white opaque or transparent quartz. Ill one place the ore is associated with stibnite, and the quartz frequently exhibts a

Sep 1, 1905

The Candelaria deposit was first discovered by Spanish prospectors in 1863; but its silver ore was not exploited on a large scale until 1873, when the Northern Belle mine was opened. The first Northern Belle mill was built at Belleville, 13 km (8 miles) to the northwest. From 1876 to 1886 the Northern Belle paid over $1 million in dividends each year. In 1882 a 43 km (27 miles) pipeline was built to Candelaria from the White Mountains to the west to furnish water for the mills and town at Candelaria; it was 100 mm (4 in.) in diameter, hand- riveted, and made in England. Sections of this pipe are used. as claim posts today. At its terminus at Metallic City, above Candelaria, a large concrete reservoir was built and called Metallic Lake. It is being used today by Candelaria Partners. Production at Candelaria declined after the mid-1980s, revived for a short period in the early 1890s, and during the 1920s a 300 ton mill operated for three years. Candelaria Partners is owned jointly by Occidental Minerals and Congdon and Carey, with Occidental being the operator. An engineering study done in mid-1979 resulted in engineering for the plant commencing in late that year and heap leaching being started in September 1980. Two million tons of ore per year containing 108 g Ag and 0.28 g Au per mt (3.15 oz Ag and 0.008 oz Au per st) will be mined from the Lucky Hill and Mount Diablo .areas about 1.6 km (1 mile) up Pickhandle gulch from the heap leaching operation. Waste -to -ore ratio is about 2:l. Ore is mined by front-end loaders and hauled by 85 ton trucks to the primary crusher, a 1,070 by 1,780 mm (42 by 70 in. ) gyratory. Primary crushed ore is conveyed to a stockpile of 10,000 tons live capacity from which it is reclaimed by hydraulic reciprocating feeders and conveyed to the crushing and agglomeration facility. This unit is composed of a 1.8 m by 2.6. m (6 ft by 12 ft) double deck screen, a 2.1 m (7 ft) standard cone crusher, and three conveyors. At the conveyor transfer points NaOH and NaCN solution is sprayed on the ore as it falls through a row of 75 mm (3 in.) hanging steel bars. During this process of turning and dividing the ore, the fines agglomerate with the coarse ore to give better percolation for the leach solution on the heaps. Provision was made for the later addition of 2 kg per mt (4 lb per st) of cement to aid in the agglomeration. Agglomerated ore at minus 50 mm (2 in.) size is then conveyed to a 390 mt (430 st) loadout bin where it is transferred to 85 ton rear dump trucks for haulage to the leach pads.

Jan 1, 1981