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By H. E. Criner, M. G. Driessen

THE cyclone thickener has two important applications in wet washing plants: (1) water clarification, and (2) fine coal recovery. The thickener consists of a conical chamber into which the fluid is injected through a tangential nozzle at the periphery of the large end. The solids of the slurry are concentrated and removed from the apex of the cone while the "thinner" portion is discharged from a central tube in a flat head at the large end, fig. 1. The advantages of the cyclone thickener are that large volumes of slurry may be treated with small space requirements and that the underflow concentration may be controlled easily and readily to meet varying feed conditions. Control of Washery Water Solids Concentration: The first requirement, in the process of controlling solids concentration, is that the solids be removed from the body of circulating water at the rate at which they are introduced. If this condition is satisfied the concentration will remain at a fixed value and will not increase to uncontrollable levels. The second requirement is that the solids concentration

Jan 1, 1951

By Ferid Kromer

Geography: The Tagtepe district of the Vilayet of Eskigehir is about 20 miles northeast of the city of Eskigehir (approximately midway between Ankara and Istanbul) in western Anatolia. The area is a mountainous one, the highest peak being Tagtepe Mountain (5200 ft) which is approximately in the center of the district. The mountains drop off to the deep valley of the Sakarya River on the north and to the plain of Eskigehir on the south. FERID KROMER, Junior Member AZME, is a Consulting Mining Engineer and General Manager, Bagtag Turk Maadin Ltd., Istanbul, Turkey. New York Meeting, February 1950. TP 2629 H. Discussion of this paper (2 copies) may be sent to Transactions AIME before Feb. 28, 1950. Manuscript received Dec. 29, 1948. For the most part, the watershed is on the northern side of the mountain barrier, draining into the Sakarya River, which in turn empties into the Black Sea midway between Zonguldak and the mouth of the Bosphorus. The approximate area covered by the Tagtepe district is shown in fig. 1. Transportation to shipping points is available via the Istanbul-Ankara railroad. The station on this line nearest the mining district is Alpikoy station, about 20 miles by road southwest of Tagtepe Mountain. Interior roads within the district are poor. Being of dirt, the winter rains and snows render them almost impassable for trucks from about the middle of December until the end of March, thus presenting a considerable transportation problem. However, the roads from the Bagoren and Tagtepe chromite mines to the railroad shipping point at Alpikoy station have recently been repaired and will be maintained for all-weather truck transportation. Detailed climatic data are not available. However, in general the spring, summer, and early autumn months are dry, and good weather may be expected from May until early November. Then the winter rains commence, and heavy snow is usual during January and February. Geology: The mountainous structure of Tagtepe belongs to basic rocks of serpentine (Variscan Orogeny) which is in contact with Paleozoic schists at west, and an Oligocene outcrop of red clays in Margi-Sepetci region (see fig. 1) at southeast. The northwest and southwest borders of Tagtepe district are, respectively, surrounded by Paleozoic schists and pebbly gray and yellow Neogene clays. More recent formations of alluviums overlay the plain of Eskigehir. Dark basic rocks of trachytes with hornblende are visible on Turkmenbaba Mountain, at the west of Tagtepe. Mineral Occurrences: Chromite: the most important mineral found in any quantity in the Tagtepe district. The alignment of the deposits of chromite is in general along the line Bagoren-Tagtepe (see fig. 1). The first mines in the area were those of Tagtepe and Bagoren, which were developed over 20 years ago with Swedish capital. Other deposits of chromite, more recently discovered and so far of less importance, are being worked at Kurucor, Komurcu, Gelinmezari, and Lacin (see fig. 1). Deposits average generally between 46 and 48 pct chromic oxide, with the exception of the Bagoren mine which averages 44 pct. However, a new lode, very recently plotted, in the Tagtepe mine averages 50 pct Cr,O,, 4.6 pct SOz, and 7 pct FeO. Geological character of the chromite occurrences in the Tagtepe mine may be considered typical of most chromite lodes in this area. The indications are that the formations of ore lenses are developed by the segregation of chromite crystals intruded into the serpentinized rock, and exposed later to tectonic movement within the zone of crystallization. All lenticular masses are more or less regular in shape and follow each other in southeast-northwest direction and dip generally 70" NE. Ore lenses do not seem to persist in depth, average depth of two lenses is 60 ft below surface. Three lodes .have been mined as open-pit. The average dimensions of individual lenses are as follows: pitch length, 100 ft; breadth, 27 ft; and width, 20 ft. The lenses and their enclosing rocks are broken by parallel fractures in approximately east-west direction. These joints are filled, except in one lode, with cementing material, which gives to the ore a

Jan 1, 1951

By Herbert A. Stein, John B. Murdock

AN analysis of perlite expansion furnaces must be based upon one consistent theory which explains how and why perlite does expand when heated. There is more than one such theory, so to establish a basis for the rest of this paper, the authors will outline one theory and the data that support it before going into the construction details of different furnaces. JOHN B. MURDOCK, Associate Member AIME, is President, and HERBERT A. STEIN is Secretary, of the Perlite Corp., Phoenix, Arizona. San Francisco Meeting, February 1949. TP 2736 H. Discussion of this paper (2 copies) may be sent to Transactions AIME before Feb. 28, 1950. Manuscript received Jan. 3, 1949. Why Perlite Expands: Perlite is known to contain some dissolved or combined fluid, probably water, which is lost when the material is heated to fusion. It is generally agreed that when particles of this volcanic glass are heated to fusion in such a way that vaporization of this water into steam takes place at the same time as the glass becomes soft, the steam puffs up each particle into a mass of glass foam. It is a simple matter to expand perlite under a blow torch for demonstration purposes, but in commercial installations, the problem is made more complicated by the necessity of making efficient use of both raw material and fuel, and producing an expanded aggregate conforming to close specifications for size distribution and density, using many different types of perlite rock. The ability of a given operation to fulfill these requirements depends upon the proper handling of four variables. They are: (1) the nature of the perlite rock used, (2) the heating time taken to get the perlite particle up to its highest temperature, (3) the size of the particle, and (4) the maximum temperature the particle reaches. Types of Perlite Rock: Perlite may be classified in many ways, but the characteristics most easily correlated with its behavior in furnaces are its softening temperature, its effective water content, and its "liveliness." Actually there is a close correlation among these three characteristics so it is possible to describe a perlite using any one of them. Lively perlites always have low softening temperatures and high effective water contents, whereas the deader perlites have higher softening temperatures and lower effective water contents. These softening temperatures run from 1400°F to above 2000°F and the effective water contents run from over 2 pct down to 0.2 pct. Included in the category of lively glasses are the green pitchstone from the Edgar claim in Superior, Ariz., the red pitchstone on the Animas highway south of Lordsburg, N. Mex., and a gray glass from Beatty, Nevada. Examples of intermediate perlites (in order of decreasing liveliness) are the Guzman-Lobb perlite with obsidian

Jan 1, 1951

By J. P. Smith

DURING the winter of 1946 to 1947 potash operations in Germany, France and Spain were visited by the author. The U. S. Department of Commerce, through its Field Intelligence Agency Technical, sponsored the inspection of German properties by representatives of the American Potash Industry. Vast deposits of potash salts occur in the above-mentioned countries, and the production and marketing of these salts are important in their respective economies. Germany An area of some ten thousand square miles of north central Germany is underlain by lenses of potash salts interbedded with greater thicknesses of sodium chloride or common salt. These salt measures are relict deposits from the evaporation of an arm of the Zechstein (Permian) Sea which formerly covered northern Europe. During the peak year of 1940, Germany produced salts containing 1,860,000 tons of K,O and, next to coal, potash is its most important mineral raw material. The German deposits are found in the synclinal areas flanking the Flechtingen, Hartz Mountain and Thuringian uplifts, and, to the north they lie beneath the plains of Hannover. The regional dip of the salt measures is to the north, hence, lenses of potash found at depths of 500 to 1000 m* in the Two general depositional provinces of potash concentrations are recognized: (1) the main basin, and (2) the subsidiary basin. The main basin embraces the producing districts of Stassfurt, Sud Hartz, Halle, and Hannover-Braunschweig. From J. P. SMITH, Member AIME, is Clzief Geologist, U. S. Potash Co., Carlsbad, New Mexico. New York Meeting, February 1948. TP 2621 H. Discussion of this paper (2 copies) may be sent to Transactions AIME before Feb. 28, 1950. Manuscript received May 26, 1948. 40 to 45 pct of the German production comes from this basin. The Werra-Fulda district, producer of 50 pct or more of the German potash, lies in the subsidiary basin. Salt Stratigraphy: Five commercial potash zones or beds are recognized. The three beds present in the main basin are: The Stassfurt bed that occurs near the top of the older salt, and the Ronnenberg and Riedel beds in the younger salt occurring 200 to 300 m, respectively, above the Stassfurt bed. The Stassfurt bed is found over the entire main basin, while the Ronnenberg and Riedel are restricted to the Hannover area. Two potash beds occur in the Werra basin. These are the Thuringian, or lower bed and the Hessen or upper potash bed that occurs 60 m above. These beds lie 210 and 150 m, respectively, below the position of the Stassfurt bed, and are restricted to the Werra basin. The upper bed is 2.5 to 11.5 m in thickness, and the principal ores are hartsalz with varying thicknesses of carnallite. The lower bed averages 3 m in thickness, and contains hartsalz. The upper bed is presently being mined, and the

Jan 1, 1951

By Lauren A. Wright

SINCE the early nineteen-thirties the production of talc* in California has increased five-fold to a yield in 1947 of about 76,000 tons (fig. 1); conse- .' Unless otherwise qualified, the term "talc" in this paper will imply a mixture of the pure talc with such related minerals as commonly occur in the commercial material. quently the state is now one of the nation's three principal talc producers. The 1947 output, which represented about one-sixth of the national total for that year, was approximately equal to production from Vermont, and was second in value to that from New York. New York has an annual output of about 130,000 tons. Three factors have been mainly responsible for the expansion of talc production in California. (1) Technical advances and innovations, particularly in the ceramic industry, have increased the demand for certain California talcs. The nation's high-qual- LAUREN A. WRIGHT, Junior Member AIME, is Associate Geologist, California Division of Mines, San Francisco, Calif. Los Angeles Meeting, October 1948. San Francisco Meeting, February 1949. TP 2622 H. Discussion of this paper (2 copies) may be sent to Transactions AIME before Feb. 28, 1950. Manuscripts received Dec. 7, 1948; revision received Aug. 29, 1949. This paper is Contribution No. 508 of the Division of Geological Sciences, California Institute of Technology, Pasadena, Calif. the industry, mainly through demand for the urgently needed steatite. Fibrous talc, used as a paint extender by the armed forces, also was produced in quantity. (3) Expanding western industry has continued to draw heavily on California talc deposits, and has extended the upward production trend into the postwar period. Considerably more than half of California's current talc output is used as ceramic raw material. The paint, rubber, paper, and textile industries are the more important consumers of the remainder. Smaller amounts are used as insecticide carriers, and in cosmetics, and various minor industrial applications. During World War I1 the principal steatite deposits were studied in detail by members of the U. S. Geological Survey; other sources of commercial talc in the state have in the past received but scant geologic attention. The following discussion is essentially a preliminary report based on research currently in progress by the California Division of Mines, and includes a brief outline of previous steatite studies by the U. S. Geological Survey. Distribution of California Talc Deposits: The principal talc resources of both California and Nevada are confined to an elongate belt that extends from the vicinity of Baker, in north-central San Bernardino County, northwestward to include the southern Death Valley region, the Inyo Range, and the Palmetto-Oasis district of western Nevada (fig. 2). This belt, of which all but the northernmost tip lies in California, is approximately 200 miles long and 30 miles in average width. Included in the belt are four contrasting areas. The deposits within each area are geologically similar and yield talcs that have characteristic properties and applications. Three of these local "talc provinces" lie within California's portion of the talc-

Jan 1, 1951

By C. Kremer Bain

UNDER present economic conditions the necessity of mechanizing the mines of our country has become a very important problem. More and more mines are looking toward increased or complete mechanization of their operations, so in the writer's opinion it is quite pertinent that some of the associated problems that accompany such a program be pointed out. The following generalization on attempting a comparatively slight change in mining methods at the Southeast Missouri properties of the St. Joseph Lead Co. is an attempt to point out to anyone faced with a mechanization program that much time and C. KREMER BAIN, Member AIME, is Assistant to General Manager, St. Joseph Lead Co., Bonne Terre, Mo. AIME Sun Francisco Meeting, February 1949. TP 2743 A. Discussion (2 copies) may be sent to Transactions AIME before March 31, 1950. Manuscript received Dec. 27, 1948; revision received Oct. 13, 1949. effort can be saved, and many headaches avoided, if company policy can be established whereby everyone in the organization from the top down is desirous of improvement and is working towards the common objective. Any changes in mining methods, machinery, and so on, must entah coordination of all the other phases of mining as they all have to mesh as accurately as a smooth running chain of gears if satisfactory results are to be obtained. The purpose of this paper is to point out some of the problems which :have to be surmounted when any mechanization program is attempted, rather than to go into the mechanical details and actual operating data of the equipment employed. After all, machines have to be adapted to best suit the local conditions, so no general formula can be given that would apply to other properties. Although St. Josepll Lead Co. has been mechani-cally-minded for a number of years, as shown by the fact that over 110 million tons have been loaded mechanically since 1922, any further mechanization or changes always present a problem. When our ore bodies were of great height and area the breaking and loading were developed to a high degree of efficiency according to known standards. But in recent years, when our ore bodies became smaller and more scattered, the amount of tonnage we were able to break was not sufficient to supply our shovels with enough rock to keep them working to capacity. Our first step in attempting to rectify this situation was to develop a technique of moving the shovels to the broken ore in various stopes, which sometimes entailed a move each shift. A shovel is not producing when being moved, so within the last two years we have been endeavoring to devise some means of getting more rock broken in these smaller stopes and it was then that we became interested in jumbo drilling. To illustrate the fact that mechanization is a process of a chain of developments rather than a stroke of genius can best be shown by a general summary of conditions and obstacles which were encountered in our present attempt to establish template drilling of a standardized round. Our first jumbo was more or less the conventional type with two machines mounted on a crossbar. This arrangement proved to be impractical to duplicate satisfactorily the regular routine of slabbing in a more or less radial pattern as carried out by column drilling. One machine was either too close to or too far from the face and there was another disadvantage in that one drillman was frequently delayed while waiting for his partner to bottom a hole. Our next step was to eliminate the crossbar so the machines could operate independently of each other, but we still encountered the same objections mentioned above except that one drillman was free to drill three holes instead of one in a vertical plane independently of his partner. Time studies while trying to adapt the jumbo to the regular slabbing practice as accomplished by column drilling, revealed that twice as much time was spent in maneuvering the jumbo into position and aligning the holes as was consumed in drilling.

Jan 1, 1951

By S. S. Arentz

PLANNED health and safety programs have become an essential part of American industry because such programs lead to increased operating efficiency, improved labor relations, better public relations, and to substantial savings in compensation insurance. Those of you who have had the unpleasant duty of informing the wife or widow of one of your men of his serious injury or death while on the job, know that all the benefits of a successful safety program do not show on the balance S. S. ARENTZ, Member AIME, is General Superintendent, Nevada Operations, Combined Metals Reduction Co., Pioche, Nevada. AIME San Francisco Meeting, February 1949. TP 2741 A. Discussion of this paper (2 copies) may be sent to Transactions AIME before March 31, 1950. Manuscript received Jan. 6, 1949. sheet. These programs are of particular importance to the mining ,industry because mining's reputation as an unusually hazardous industry and the commonly isolated location of mining operations tend to focus attention on these problems. Description of Operations: Before proceeding with a discussion of our health and safety programs at Pioche, it may be proper to give a brief description of Pioche and of our operations there. Pioche is one of the early Nevada mining camps. It was founded shortly after the discovery of high grade silver ore in 1863 and mining has continued with more or less regularity to the present day. In an era of lawlessness, Pioche was notorious. The story persists that 75 men died with their boots on before one died a natural death, and old payroll records show that nearly as many gunmen were employed to stand off claim jumpers as there were miners working the mine. That was probably as close to a safety program as the times permitted. Pioche is situated in southeastern Nevada on the main highway between Ely and Las Vegas. The camp is on the flank of "Treasure Hill," near the original silver discovery, at an elevation of about 6000 ft. The present day population of about 2000 is primarily dependent upon the mines of the area, although Pioche also serves as the county seat of Lincoln Couqty and as the center of the surrounding livestock industry. The camp is served by a branch of the Union Pacific Railroad and receives power from the generators at Hoover Dam. The Pioche operations of the Combined Metals Reduction Co. were started in 1923 when the first complex lead-zinc ore was shipped to the company's mill at Bauer, Utah. The modern mill at Pioche was completed in 1941. The operations are medium sized in the nonferrous field, employing an average of 350 men in the mine, mill, and related works. The complex lead-zinc ore is mined from replacement deposits in a comparatively flat, extensively faulted, limestone horizon. Mining methods vary from stull-supported open stopes to filled square-set stopes. The thin bedded limestone and shale overlying the ore is allowed to cave as areas are mined out and caving frequently follows closely upon ore extraction. The relatively heavy ground and the numerous faults add to the problems of safe mining. The mine is well mechanized and the mill and surface plant are modern and well equipped. Labor is organized in a C.I.O. union and labor-management relations have been unusually harmonious. During most of the period since 1923 a competent supervisory staff worked to reduce safety hazards but the primary responsibility for safety rested on the individual workman. Accidents happened and all too frequently they were regarded by all concerned as unavoidable. In October 1939, the late Robert L. Dean became superintendent at Pioche. Most of his previous experience had been in the fields of iron and coal mining and from that experience he brought the concept that no accident is unavoidable. Many of the features of our present health and safety programs were initiated by Mr. Dean during his term as superintendent. Health Program: Our health program centers in Dr. Q. E. Fortier and his new, well-equipped, and well-staffed, modern hospital in Pioche. The program starts with a thorough pre-employment physical examination and is followed by yearly re-examinations at the expense of the company. The Pioche Mutual Benefit Association, to which all Pioche mine operators and employees belong, pays benefits covering hospitalization and surgery expense incurred by employee members and their families. The Association is governed by a board of directors elected by its members. The mine operators of the district donated the original capital and pay the monthly dues of the employee members. The employees pay the dues covering members of their families. Though not strictly a part of the

Jan 1, 1951

By Richard Hamburger

(1) Introduction: The purpose of this paper is to call attention to certain methods which will shorten and simplify the calculation of plane tri-angulation. These methods, though not new,',2 do not appear to have been properly publicized in the United States. The cotangent (or tangent) solution of a triangle permits calculation of coordinates in a smooth series of operations of the calculating machine with a minimum of notation. The semigraphical solution permits the simultaneous solution of a number of intersections, at one station, without the added burden of solving additional triangles. This in turn encourages more numerous field checks and more complete office computation than would normally be undertaken. RICHARD HAMBURGER, Junior Member AIME, is Junior Engineer, Inspiration Consolidated Copper Co., Inspiration, Arizona. TP 2798 A. Discussion (2 copies) may be sent to Transactions AIME before March 31, 1950. Tables I and I1 show the distinct advantage which the combined use of the cotangent and the semi-graphical solutions hold over, what might well be called, the standard solution. It will be shown that the positions for the entire net in fig. 1 may be calculated without computing any distances, thus saving much computation. (2) Comparison of Cotangent and Standard Solutions: To compare the cotangent method with the standard method, the notation involved in the solution of triangle 1-X-4, fig. 1, by both methods is shown. Fig. 2 shows the given information, that is, the coordinates and azimuths of Stations 1 and 4, and angles of the triangle or the azimuths from Stations 1 and 4 to Station X. (Azimuths are used here because the author feels that they are easier to handle, but bearings may be used by those who prefer them). It will be seen from the above comparison that the advantages of the cotangent solution of a triangle are: (a) rapidity, (b) full use of the calculating machine, (c) the small number of

Jan 1, 1951

By E. H. Crabtree, V. W. Winters, T. G. Chapman

IN 1939 one of the authors* described advances in carbon-cyanidation for the period 1932 to 1939 • T. G. Chapman: A Cyanide Process Based on the Simultaneous Dissolution and Adsorption of Gold. Trans. AIME (1939) 134, 207. and included: (1) the dissolution of gold in an ore pulp by cyanide and its simultaneous adsorption by carbon, (2) the stage addition of adsorptive carbon and its movement countercurrent to the flow of the ore pulp, and (3) the recycling of adsorptive carbon. Considerable additional experimental and pilot plant work have been done by the authors since 1939 and the results of such work E. H. CRABTREE, JR., Member AIME, is Director of Milling, Eagle-Picher Mining and Smelting Co., Miami, Okla., V. W. WINTERS is Superintendent, Sahuarita Concentrator, Eagle-Picher Mining and Smelting Co., Tucson, Ariz., and T. G. CHAPMAN, Member AIME, is Dean, College of Mines, University of Arizona, Tucson. AIME Sun Francisco Meeting, February 1949. TP 2744 B. Discussion (2 copies) may be sent to Transactions AIME before March 31, 1950. Manuscript received March 3, 1949. have been sufficiently encouraging to justify a second paper. Developments from 1939 to 1948: In 1941, Verne W. Winters did the first experimental work in eliminating agitation during dissolution and adsorption and thereby demonstrated that the dissolution of gold by cyanide and its simultaneous adsorption by carbon take place essentially to completion in a stagnant pulp. Shortly thereafter, and as a result of the stagnant pulp experimental work, Winters and Chapman demonstrated that soluble gold, in a cyanide ore pulp, would migrate from the ore pulp through a perforated container and be adsorbed by carbon within the container. During the same year, the Eagle-Picher Mining and Smelting Co. erected a 25-ton capacity pilot plant at the Harquahala mine about 9 miles from Salome, Ariz. Experimental work at this plant comprised chiefly the stagnant pulp modification of carbon-cyanidation although considerable work was done in the grinding and activation of carbon, and testing various methods of extracting adsorbed gold from carbon. Soon after the completion of the Harquahala

Jan 1, 1951

By T. B. Counselman

"C" VERYONE who has to calculate cyanide circuits, -¦-' using either thickeners, filters or both, realizes the headaches involved in solving a set of simultaneous equations. When you calculate a large number of these, based on experimental data, trying to arrive at the best flowsheet, you can spend days at the task. T. B. COUNSELMAN, Member AIME, is Manager, FluoSolids Division, The Dorr CO., Stamford, Conn. AIME New York Meeting, Feb. 1950. TP 2783 B. Discussion (2 copies) may be sent to Transactions AIME before March 31, 1950. Manuscript received Aug. 23, 1949. Herewith is presented a short method, equally accurate with the standard simultaneous equation method, and much quicker. Here for instance is the standard simultaneous equation method, for 100 tons per day of $10 ore, with the conditions assumed as below (fig. 1). Conditions Assumed a. 100 tons of ore per day crushed in cyanide solution. b. Discharge from all thickeners with 50 pct moisture. c. $10 value dissolved per ton of ore. d. 50 pct in mill and 50 pct in agitators. e. 400 tons of solution from thickener V precipitated to $0.02. f. Agitation with a dilution of 2 of solution to 1 of solids. g. Let V, W, X, Y and Z represent value in dollars per ton of solution discharged from the resvective thickeners. Equating Out of and Into Each Thickener 1. lOOVplus 400V = 500W plus (0.50 X $10 X 100) 2. 1OOWplus 600W= 500X plus lOOW plus (0.50 X $10 X 100) plus lOOV 3. lOOXplus 500X = 1OOWplus 500Y 4. 1OOY plus 500Y = 1002 plus 1OOX plus (400 X 0.02) 5. 100 Zplus l00Z = l00Y plus 100 tons of water (value $0.00) Simplifying: 1. v = W plus 1.00 2. W = X plus 1.20 3. X = Y plus 0.24 4. Y = 0.2Zvlus 0.064 5. 2Z= Y

Jan 1, 1951

By S. Korman, S. B. Tuwiner

THE purpose of this study is to investigate the effect of agitation upon the flotation of chal-cocite. It is believed that the concentration of collector which is required to produce a water-repellent surface upon the mineral is determined among other things by the rate of collector diffusion through an impeding film of more or less stagnant fluid. Its thickness depends upon the intensity of agitation. In the first part of this study a method is developed for determining the thickness of the fluid film upon a polished surface of synthetic chalcocite. The film thickness is then determined as precisely as possible under three readily reproducible condi-tions of agitation intensity. These determinations serve as a quantitative evaluation of agitation. In Part 2 it is demonstrated that in solutions of lime at PH 10.5 and in the absence of air or dissolved oxygen the concentration of collector required to develop water-repellency upon the surface in any given time is a function of the fluid film thickness, and that after about 10 min the relation becomes linear. This points to the possibility that the collector ions may be involved in a diffusional process through the fluid film. In Part 3 it is shown that where air is present more collector is required to produce water-repellency than for the case in which air is excluded. Next, it is shown that for the latter case the most minute amount of sodium sulphide suffices to preclude development of water-repellency. Both oxygen and sulphide are in a sense, therefore, depres- sants for chalcocite in a lime solution at PH 10.5. When oxygen and sulphide are present simultaneously there is a region of minimum critical collector concentration for water-repellency which is very markedly lower than the concentration where oxygen alone or where sulphide alone is present. Moreover, this minimum appears to be identical with the collector concentration which is required where sulphide and oxygen are both excluded. This would appear to indicate that in certain ranges of concentration of oxygen and sulphide, each cancels the effect of the other. The range of concentration of solution for this effect in solutions saturated with oxygen is a function of the agitation intensity. Thus we are led to believe that a chemical reaction involving sulphide and oxygen occurs at the surface of the chalcocite. Part 4 of this study is a further development of this idea and some of its consequences from an electrochemical standpoint, while in Part 5 the disappearance of sulphide from pulps of several minerals is noted with an indication that both dissolved oxygen and the mineral are necessary, the reaction being catalytic or pseudocatalytic. Part 1. Determination of Fluid Film Thickness The Fluid Film Concept: The hindrance to the diffusion of collector to the surface of the chalcocite is a barrier of stagnant fluid in the vicinity of the surface. It is the same barrier that limits the rate of transfer of heat, the rate of solution of dissolving crystals, or any of the other rates pertaining to chemical engineering unit operations in which the film resistance to diffusion plays a principal role.' Fluid films are present upon the surface to an extent which varies with the geometry and intensity of agitation. They result from the fact that in the region within the fluid near a surface, motion is entirely parallel to it. From the standpoint of transfer to or away from the interface the effect is as though the fluid were stagnant. Outside of the film, movement of the fluid may be streamline or turbu-

Jan 1, 1951

By S. N. Kesten, Richard Murphy, A. H. Land, W. F. James

Introduction—by W. F. James and A. H. Lang This paper describes the geology and present state of development of uranium and thorium deposits in Canada. It is expected that this information will be added to that now being assembled from available American and British data to provide a background for engineers and others who may be called upon in the future to examine or search for radioactive deposits. Some generaliza- tions regarding the Canadian deposits are included, but many of these are necessarily tentative. The information contained in the paper is up to date for the end of 1948. The information contained in this paper has been derived from many sources. Much of it is from the records of the Canadian government-owned Eldorado Mining and Refining (1944) Limited and from examinations of government and private properties made by officers of the Geological Survey of Canada. Some information on privately-owned properties has been quoted from other sources mentioned in the text. The information has been assembled by W. F. James of Eldorado Mining and Refining (1944) Limited and A. H. Lang of the Geological Survey of Canada, who have written the sections for which they are credited. The description of the Eldorado mine is based on reports by Richard Murphy, formerly chief geologist for Eldorado; the report by Mr. Murphy was published previously by the Canadian Institute of Mining and Metallurgy,' to whom the writers are indebted for permission to repub-lish. The section dealing with the geology of and development work in the Goldfields area, where Eldorado has important holdings, is by S. N. Kesten, field geologist for Eldorado. The geological diagram of the Goldfields area (fig. 2) was kindly supplied by A. M. Christie of the Geological Survey of Can-

Jan 1, 1951

By W. W. Anderson

WEBSTER'S dictionary gives the following definition for "efficiency": "Effective operation as measured by a comparison of actual and possible results." Engineers think of this definition in terms of the ratio of output to input, and since they learn early in their careers that output cannot be greater than input, they usually consider a machine to be 100 pct efficient if it operates perfectly, and to be 0 pct efficient if it performs no function at all. Hence, values of efficiency between 0 and 100 pct reflect how well or how poorly a machine has performed; and any method of calculation for determining efficiency that does not show 0 pct when no function has been performed, and 100 pct when a perfect job has been done, is looked upon with skepticism. The literature on performance of coal cleaning equipment contains many references to efficiency of performance, but there is no unanimous agreement among the authors as to how the efficiency should be determined. Numerous methods of calculation have been proposed from time to time, most of which attempt to fulfill the definition of efficiency by comparing output to input, using float and sink data or some other characteristic, such as dry ash content, as a measure. However, to the best of our knowledge, none of the methods advocated in the past meets the requirements of a practical engineer in judging cleaning plant performance, because the formulas previously advocated do not confine numerical values of efficiency between 0 and 100 pct, and the literature does not contain an explanation of the meaning of the formulas under various conditions of operation. Therefore, the following formula, based on Yancey's concept of misplaced material,' is proposed as a measure of efficiency which will conform with the requirement that the values of efficiency must be confined between 0 and 100 pct: Quantitative Efficiency = 100.0 — (pct wt sink* in cleaned coal + pct wt float* in refuse)t • At gravity of separation effected by cleaning equipment. t Total product basis of the size of material under study. It will be noted from this formula that an actual separation between cleaned coal and refuse must be effected by the cleaning equipment before the formula can be used; also, that improperly distributed material in the products must be determined at the gravity of separation, rather than at some arbitrary test gravity. It will further be noted from this formula that the weight of cleaned coal recovered and the weight of refuse rejected by the equipment have their proper influence, because the sink in the cleaned coal and the float in the refuse are expressed in the equation as a per cent weight of the total products of a definite size discharged from the plant. This formula has one other important characteristic, in that it requires no knowledge of the feed to the cleaning unit. Comparison of data on the products from a cleaning unit with the feed to the unit can lead to many worthwhile conclusions. but not in regard to the efficiency of separation effected by the equipment. The reason for this is that mechanical fracture of many pieces of the feed occurs during the cleaning process; and in some processes crushers are deliberately incorporated in the flow to produce size degradation to free impurities from cleaned coal, thereby altering the washability characteristics of the material while in process. When such size degradation occurs in a cleaning process, it is quite possible that this crushing will result in a reduction in ash content in the finer sizes of cleaned coal when compared with the raw coal feed, even though no true separation has been effected. This is due to the fact that the finer sizes have been "sweetened" by the crushing of cleaned coal from some coarser size. Many formulas advocated in the past show high cleaning efficiencies

Jan 1, 1951

By W. W. Anderson

John Griffen—The author has called attention to an important phase of coal cleaning since there has been considerable loose thinking regarding efficiency formulae and particularly inaccurate application of ones that have been advanced. The unanimous acceptance of an efficiency formula has been prevented because the issue has been beclouded by the fact that the objective of cleaning—-the control of chemical characteristics of the coal such as ash or sulphur content—is attained by an indirect rather than a direct method. The indirect method is usually to effect separation according to differences in specific gravity. Such separation relies on the fact that the desired chemical qualities are roughly inversely proportional to the specific gravity of the particles being separated. However, the sulphur content of the materials found in raw coal usually exhibit a far from straight line relationship with their specific gravities. Mr. Anderson meets this issue squarely by basing his formula on the efficiency of separation according to specific gravity only. It is true that machines that effect primarily a specific gravity separation are in no way directly influenced by the chemical analysis of the materials being handled and it can be argued that it is illogical to measure their efficiency by any criterion other than that of separation according to specific gravity. For the moment, let us accept this basis and analyze the application of the efficiency conception and formula proposed by Mr. Anderson under conditions which he has not considered but that will readily be found. Let us take the separation shown in his table III as an example. He terms this an example of a high order of quantitative efficiency. We will show the weight distribution of the various gravity fractions as between clean coal and refuse in table XV. Only in the gravity range of 1.50 to 1.70 has material been misplaced and, on the assumption that the cleaned coal and refuse are equal in weight, Anderson shows a quantitative efficiency of 99 pet. The gravity of separation is 1.60 and one third of the 1.50 to 1.60 sp gr fraction is misplaced in the refuse while one third of the 1.60 to 1.70 sp gr fraction is misplaced in the cleaned coal. Let us now see what effect the gravity distribution of the total products has on the quantitative efficiency calculated by his formula. Still using the data in table

Jan 1, 1951

By H. A. Baumann, C. C. Wright, T. S. Spicer

A LTHOUGH the efficiency of coal utilization has A improved steadily since the turn of the century as a result of continued research and development, little of this increase in efficiency can be credited to improvements in the single retort stoker. There are still numerous problems connected with the burning of coal on this type of unit, the solution of which would be a stimulus to coal and stoker sales and a boon to the user. Among the more critical of these problems is that of burning slack bituminous coal of strong caking tendency and containing large percentages of fine material. With the increased trend toward mechanical mining, it may be expected that in the future an even larger percentage of the coal available for use on this class of stokers will contain higher percentages of the fine sizes, thus aggravating an already serious problem. During the war, the problem of efficient and economic burning of available coals on the single retort industrial stoker, particularly the class 4 stokers (301 to 1200 lb per hr), became quite critical in many localities. In addition to the inherent problems associated with the use of this class stoker, the coals available were frequently not those best suited for use on this equipment and the manpower shortage made it difficult as well as expensive for the consumer to provide the labor necessary to keep the iires in reasonably satisfactory condition. Although exact figures on the number of class 4 stokers are not available, the yearly tonnages of coal consumed are of considerable consequence and have been estimated* at between 15 and 20 million the stoker research staff of the Mineral Industries Experiment Station of The Pennsylvania State College was requested to study the performance of the single retort industrial stoker and to undertake a program of research on the more efficient burning of bituminous coal on this class of unit. Various aspects of overall problem have been investigated since the inception of this program in 1942, only one of which is here reported. General Considerations There are numerous factors which appear directly or indirectly to affect the overall performance of bituminous coal underfeed stokers, such as the caking or matting properties of the coal, the contact relationships between coal and air, the ash fusion characteristics, the burning rate, and the design of the stoker and furnace in which the coal is burned. An analysis of the influence of each of

Jan 1, 1951

By Gilbert H. Cady

HE application of geology to problems arising A in coal mining engineering and coal preparation has, in general, been somewhat remote, or the geological problems have involved such simple forms of geological knowledge that the attention of specialists in this field has not been commonly thought necessary. Furthermore, the geology of coal beds and coal mining has received but scant attention in our educational system, even in mining education; so that coal mining engineers have become somewhat conditioned against geological thinking. Geologists, on the other hand, have commonly failed to realize that there are geological aspects to coal mining and coal preparation problems. A further reason why geology receives so little consideration in coal mining undertakings in this country lies in the fact that our coal supplies are, or have been thought to be, almost unlimited, have been easily accessible, and very cheaply mined. Conditions have been changing within recent years. Failures can be less easily endured. Great inroads have been made on our bonanza deposits of easily mined coal and the increasing cost of coal and the increasing difficulty of finding it in easily accessible and large quantities places greater and greater importance on accurate information and good technique in exploration. It seems probable, therefore, that the opinions and aid of geologists will be more commonly sought than in the past in the exploration for coal and in the development and operation of coal mines, and analyses of mining conditions, preparation, and utilization.

Jan 1, 1951

By J. F. Poland

Location of Basins and Geologic Features of Occurrence: The major ground-water resources of California occur and are stored in the many large alluvium-filled valleys of the state. The deposits of Quaternary age which hold the ground water commonly are called valley fill. Meinzerl has discussed the characteristic features of this fill and showed that for western United States its distribution is concentrated chiefly in western Utah, Nevada, central Oregon, California, southern Arizona, and central New Mexico (see fig. 1). For California the principal ground-water basins are shown on fig. 2. Except for the Mojave Desert and the "Basin and Range" Valleys (fig. 2, Nos. 23 and 23A), which are not shown individually, the basins outlined on fig. 2 contain all the large reservoirs in valley fill within the state. The Sacramento and San Joaquin Valleys (fig. 2) together comprise the Great Central Valley of California, which has an area of about 18,000 square miles. The Sacramento Valley contains about 5000 square miles of valley lands, approxi- mately equal to the area of Connecticut; the San Joaquin Valley embraces about 13,000 square miles, equivalent to the areas of Massachusetts and Connecticut combined. Both the area and the ground-water use in the Central Valley is far greater than for all the other shaded alluvial basins of fig. 2 added together. Except for the Sacramento Valley, all the major ground-water basins of the state are south of Sacramento. The valley fill in these ground-water basins has been transported by streams and most of it has been laid down in alluvial fans, producing the typical alluvial-fan structure of aquifers radiating as tongues from a prismatic pile of coarse, permeable debris at the canyon mouths; these aquifers are encased down slope by poorly sorted materials consisting of a clayey matrix with imbedded pebbles and boulders, interbedded with flood-plain silt. In the deeper inland structural valleys, the alluvial deposits interfinger at depth in the valley bottoms with horizontally bedded lacustrine sand, silt, and clay. The San Joaquin Valley is a good example. In the coastal valleys exposed to the sea, such as the Los Angeles coastal plain, the alluvial fans pass coastward into lagoonal and thence to shallow marine deposits. In all these basins, the gravel is coarsest and most permeable near the apexes of the fans, but tongues or lenses of clean gravel and sand may extend for many miles down the fan. Thus, these water-bearing structures are in a position to absorb stream flow near the canyon mouths and to transmit it through the buried radiating conduits.

Jan 1, 1951

By Leslie C. Richards

The competitive position of producers of industrial minerals depends upon the delivered price of their product. Freight charges are a major factor in the sales to consumers. A comparison of freight rates on movements of limestone and silica rock in the Pacific Northwest shows a wide variance in mills per ton mile for like hauls, thereby favoring some localities over others. Introduction: Sooner or later the potential miner of industrial minerals reaches the conclusion that of the various factors he must consider, the marketing of his product is. the most important. The sooner he realizes this the better, as successful industrial mineral producers .will confirm. Herein lies a difference between i.ndustria1 mineral mining and metal mining. Often in the latter, technical problems of mining or concentrating are equal to or even surpass those of marketing. This is not meant to imply that there are no problems in mining and beneficiation of industrial minerals, but these usually are outweighed by the problem of marketing. Recentlv. an official of a nationally known corporation :i the elecl.ro-metallurgical industry made an industrial mineral survey in the Pacific Northwest. In the course of his investigation he interviewed producers and potential producers of the various industrial minerals scheduled for use in the plant now being constructed in this area. He stated that he was interested in only three factors: namely, (1) Can your product meet our specifications? (2) Are you able to supply our tonnage requirements? and (3) What will be the delivered price? If there are several sources of supply of sufficient grade and tonnage, as is often the case in the Northwest, then the third factor becomes the important and deciding one. While industrial minerals are usually sold f.0.b. cars with the freight being paid by the buyer, it is the delivered and riot the f.0.b. price that determines the competitive position of the producer and his industrial mineral product. Unfortunately for both the consumers and producers of industrial minerals in the Northwest, the sources are generally a considerable distance from the markets. This necessitates long rail hauls, for which transportation charges make up the largest share of the delivered cost and the delivered price of the product. Is it any wonder then, that rail rates play such an important part in the marketing of industrial minerals? In order to sell his product, the producer must be able to deliver to the market at a competitive price. It must be remembered that his largest cost factor is rail transportation, and that he is limited to markets that he can reach competitively. Since the cost of rail transportation is so important, let us compare various freight rates in the Northwest and attempt to determine their influence on marketing. A study of all rates on all industrial minerals transported in this area, is beyond the scope of this paper. For the purposes of comparison, let us consider limestone (broken, crushed, or ground), and silica rock. These are widely used industrial minerals, and are carried by several different rail routes from.the sources of supply to the centers of consumption in the Pacific Northwest. Markets and Sources of Supply: Limestone has a number of uses in western Oregon and Washington. Among them are agricultural lime for the soils of the Willamette Valley and southwestern Washington; paper rock for the mills in such cities as West Linn, Oregon City, Newberg, St. Helens, and Salem, Oregon, and Camas and Everett, Wash.; chemical and metallurgical-grade limestone for the metallurgical plants centering around Portland, and the Tacoma-Seattle areas'; and sugar rock for the sugar refineries at Toppenish, Wash., and Nyssa, Oregon. Since limestone for cement manufacture is generally processed near the quarry, transportation

Jan 1, 1951

By R. A. Kinzie

THE major operating costs in a cement plant are labor, power, and fuel. The opportunities and methods of savings in labor and power parallel other industries. Because our industry's use of fuel is large and its method of use is unique, the devices adapted for saving fuel are of interest. In normal practice the raw material in the form of a dry powder or a mud is fed to a rotating inclined cylinder where it is dried, calcined, and clinkered. In a Lepol kiln the raw material is rolled up into little balls with water in the nodulizer, fed to a traveling grate on which it is dried and partially calcined and tben to a short kiln where the process is finished. The more usual methods of saving heat have been the use of long kilns with chains in wet process plants and the use of waste heat boilers in dry plants (fig. 1). At the Santa Cruz plant we have used a Lepol system kiln and achieved a fuel 'economy of 4.7 gal per bbl of clinker or about 704,-000 Btu per bbl. This can be compared with our old kilns using the same fuel and raw material which consume 7.8 gal per bbl or 1,130,000 Btu per bbl. These are gross Btu and not the net figures usually used in reports of foreign companies. Our average consumption for 1948 was 708,000 Btu per bbl of Lepol clinker, based on cement sold, and oil purchased. The overall fuel consumption for 1948 was 6.61 gal per bbl. The Lepol system consists of a nodulizer, a grate, and a kiln. There is the usual accessory equipment which is similar to that used in other cement plants. The following description covers the equipment in operation at our plant. The nodulizer is a cylinder 8 ft 10 in. in diam by 18 ft long set on a slope of 1/2 in. per ft, and rotated at about 10 rpm. It is driven by a 50 hp wound rotor motor drawing about 9 kw. The inside of the shell is cleaned by a continuously operating scraper which draws 0.1 kw. The dry raw material is fed to the nodulizer where, being sprayed with water, it rolls into small spheres about 90 pct being between 1/2 and 1 in. in diam and containing about 15 pct water. The nodules fall on the traveling grate where the hot gas from the kiln is pulled down through the

Jan 1, 1951

By Raymond B. Ladoo, C. A. Stokes

T is unfortunate that the word "economics" has -¦¦ come to mean, in the minds of many people, a sort of half-baked mixture of New Deal philosophy and bookkeeping. It may mean anything from mine cost keeping to the theory of gold as a monetary standard. As applied to the industrial mineral industries as a whole it means to us an integrated study of all of the technical, geographical, market, financial, and other factors which determine whether or not a given enterprise has a sound chance of making a business success. Of course it is impossible to generalize too freely in this large group of industries which go to make up the nonmetallic minerals. This group includes such dissimilar products as crushed stone and ca- pacitor mica, phosphate rock and spinning fiber asbestos, roofing granules and potash salts, borate minerals and monumental granite. Yet in many respects their economic problems are similar and, at the same time, far different from those of the metals and of petroleum and coal. These basic differences are almost never recognized by people outside the industries and are rarely appreciated even by metal mining engineers. But in many instances these economic factors are fully as important as, if not more important than, technologic problems. It has been a long fight to get an intelligent and wide-spread appreciation of sound technology in the nonmetallic mineral industries. The importance of the economic side of the problem has yet to be widely understood and appreciated. The problem was outlined in considerable detail by one of the authors some 18 years ago in a long series of articles* and he has since prepared numerous talks

Jan 1, 1951