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By Roger W. Swindle

INTRODUCTION The Uravan Mineral Belt in southwestern Colorado and southeastern Utah is a uranium-vanadium mining district with a unique set of radon daughter control problems. This paper describes these problems and the methods that have been developed to deal with them. The greatest emphasis is placed on the problems caused by air stagnation because these are the most frequently encountered ventilation problems, and this is an aspect of uranium mine ventilation that is often misunderstood. VENTILATION PROBLEMS Most of the mine ventilation problems in the Uravan Mineral Belt are caused by one or more of the following conditions: (1) the highly irregular ore deposits are difficult to locate and delineate, (2) air stagnation occurs in many areas of the extremely irregular mine workings, and (3) it is often necessary to reactivate old workings with small airways that are often caved. Planning Problems Orebodies in the Uravan Mineral Belt are so discontinuous and unpredictable that they are difficult to locate and delineate. Data supplied by surface drilling and underground test holing are often misleading; very little ore may be found in the vicinity of an ore hole, and, conversely, an unmineralized hole may be adjacent to a high-grade ore deposit. Also, ore guides are often ambiguous, and their interpretation is usually a matter of opinion.

Jan 1, 1983

By W. H. Smith

IN connection with some recent work on the effect of impurities on the ductility of chromium, it appeared desirable to know the solid solubility of carbon in chromium. A literature survey indicated that this information was not available. Although considerable work has been done on the Cr-C phase diagram,&apos;.&apos; previous investigators have been more concerned with the structure and phase boundaries of the carbide phases than with the terminal solid solution. The phase diagram shown in Fig. 1 is taken from the work of Bloom and Grant&apos; and represents the most recent determination. As indicated by the dashed line, the a solid-solubility limit was not determined. Experimental Procedure Alloys of chromium were prepared from hydrogen-treated and vacuum-degassed electrolytic chromium plus spectrographic grade carbon. The oxygen and nitrogen content of the alloys was <0.002 pct. After melting, analysis of the alloys showed them to contain 0.02, 0.08, 0.15, and 0.55 pct C. Pieces of the alloys were heated in a protective atmosphere to various temperatures and then quenched after holding for a time sufficient to insure equilibrium. Microscopic examination of the as-quenched alloys for the presence of a second phase was used as a measure of the solubility limit. The heats were made in a multiple hearth arc-furnace using a zirconium-gettered static argon atmosphere. A zirconium melt was made before each Cr-C heat. Triple melting was used to insure ingot homogeneity as shown by microscopic examination. The alloys were prepared by adding portions of a 4.5 pct C-Cr master alloy to high purity chromium. The carbon contents listed previously were those obtained by analysis. The nitrogen and oxygen contents after arc melting were both <0.002 pct. Sections 1/8X1/4X1/2 in. were cut from the 100-g ingots and a hole drilled in one end in order to suspend the sample from a molybdenum wire. After the surface was carefully cleaned, a sample of each melt was hung in a mullite tube heated externally by a platinum resistance furnace connected to a vacuum system. The lower portion of the mullite tube was sealed to Pyrex and closed off several inches below the furnace. This was filled with sili-cone oil kept cold by circulating cold water around the outside of the Pyrex. Quenching into the oil bath was achieved by melting a fuse wire supporting the sample. It required about 4 sec for the sample to cool from 1400" to 600°C. This severity of quench was considered satisfactory to freeze-in the high temperature equilibrium. For tests made at temperatures of 900" to 1200°C, heating was done in vacuum; for tests above 1200°C, an argon atmosphere was used. The holding time employed ranged from 12 hr at 900°C to 6 hr at 1400°C. Experiments were performed at temperatures of 900°, 1000°, 1100°, 1200°, 1300°, and 1400°C. Microscopic examination for evidence of a second phase was done at X1500. Experimental Result The microstructures of a 0.08 pct C-Cr alloy as-cast and after quenching from 1300°C are shown in Figs. 2 and 3. A 0.15 pct C-Cr alloy quenched from 1300°C is shown in Fig. 4. The data obtained from the quenching experiments is shown graphically in Fig. 5. If the Van&apos;t Hoff equation is obeyed, a plot on a logarithmic scale of the mol fraction of solute vs the reciprocal of the absolute temperature should give a straight line. For dilute solutions the weight percentage can be substituted for the mol fraction without introducing any appreciable error. The Van&apos;t Hoff equation can then be written as where H is the heat of solution in calories per mol. The slope of the straight line on the log pct C vs 1/T plot gives the value of AH. Assuming that the Van&apos;t Hoff equation is obeyed, which is probably justified for the dilute solution of carbon in chromium, the heavy straight line shown on Fig. 5 represents the best fit of the data. This line was obtained as follows. On Fig. 5 the results of the microscopic examination of all alloys following quenching were plotted and designated as to whether one or two phases were seen. Below 1100°C all alloys showed a second phase on quenching. The heavy vertical lines shown in Fig. 5 therefore represent the possible range of the ter-

Jan 1, 1958

By AIME AIME

ALL sessions of the Mining Geology Committee at the Annual Meeting this year were held jointly with other groups, a plan that seemed to work out to the satisfaction of every one. Certain of these sessions are reported in this issue by others; the ones held on Wednesday are covered here. Ten papers, as published in the February M&M, were on the program for the Wednesday morning session, held jointly with the Society of Economic Geologists. Mr. Eckel made a brief presentation of his paper, on which there was no discussion, and Mr. Peoples presented that of which he was a joint author. D. M. Fraser then discussed a gravimeter survey for chromite in Cuba, which blocked out some 50,000 tons of ore, and F. G. Wells told of work on chromiferous sands in Oregon.

Jan 1, 1943

By J. T. Eash, E. M. Wise

MANY of the platinum metal alloys are hot-forged in the early stages of reduction from ingot and substantial quantities of platinum alloys are commercially employed at very high temperatures in ammonia oxidation units, orifices for forming glass rods and fibers and as cathodes in long-life thermionic amplifiers. They are used also for electrical heating resistors, safety blowout disks, thermocouples and crucibles. The tensile properties at elevated temperatures are important in many of these situations but data on the subject are limited. Creep data also are important, but will not be covered at this time. SCOPE AND CONCLUSIONS Tensile tests were carried out at room temperature and elevated temperatures up to 11000 C. on previously well annealed samples of platinum, palladium and several of their important alloys with compositions shown in Table 1. The results are presented graphically in Figs. 1 to 4. The tensile strengths of platinum and the platinum-base alloys (with the exception of the 10 per cent rhodium-platinum alloy) at 1100° C. are from 16.6 to 18.8 per cent of the strengths of the same alloys (well annealed) at room temperature. The strength ratio of the 10 per cent rhodium-platinum alloy is 22.3 per cent, which is appreciably higher than that of the- other materials tested. The tensile strengths of palladium and the palladium alloy at 1100° C. are from 10 to 11.2 per cent of the room-temperature values. The elongations of platinum and its alloys tend to increase at tempera¬tures above 600° to 800° C., while the two samples of pure palladium show a maximum at 600° C.

Jan 1, 1938

By H. R. Banks

The Sullivan concentrator has completed 20 years of operation. During this period a considerable amount of data has been accumulated concerning the characteristics of several types of flotation machines in their application to the problem of the differential separation of the lead sulphide, zinc sulphide and, in more recent times, iron sulphide, in the ore from the Sullivan mine. It is the author&apos;s intention to sketch briefly experiences encountered when the several types of flotation machines were introduced into the flow. Original Flotation Equipment and Ore The original flotation equipment consisted of four 18-cell lead-roughing machines, whose concentrate was cleaned in an 8-cell and recleaned in a similar 8-cell machine. Zinc concentration was carried out in four 18-cell machines. One 8-cell zinc re-treatment machine received the reground zinc middling. The 168 cells handled some 3000 tons of original daily feed. These cells were of the standard Minerals Separation type, equipped with 24-in. impellers. Spindles were driven through geared line shaft at a speed of 255 r.P.m. The impellers were the 45&apos; type, except for the last three cells of the zinc-rougher machines, which were of sub-aeration type, with full bottom shroud and half shroud for the top side of the impellers. Four-pound air was introduced. This recognition of the benefit of intensified flotation operation for "gleaning" was an important phase in the early operation and will be discussed further on in this paper. Sullivan ore during this period is described as a finely disseminated mixture of four sulphides, totaling go per cent made up of 13.6 per cent galena, 14.4 per cent mar-matite, I per cent pyrite and 61 per cent pyrrhotite. The specific gravity of the ore was approximately 4.4. There has been a gradual change in the ore content. The 10 per cent of nonsulphide constituents has increased to between IS and 20 per cent and all three sulphides have dropped somewhat. Added equipment and changing technique have made possible the treatment of 8000 tons per day of mill feed instead of the 3000 originally handled. Special Features of Flotation at Sullivan Sullivan flotation differs greatly from customary practice. The very heavy return of lead rougher middling and lead cleaner tailing, totaling at times 125 per cent of the new feed tonnage, has been the subject of much investigation. The practice is still employed. This operating condition is mentioned because it is the author&apos;s belief that the in-

Jan 1, 1947

By H. R. Banks

The Sullivan concentrator has completed 20 years of operation. During this period a considerable amount of data has been accumulated concerning the characteristics of several types of flotation machines in their application to the problem of the differential separation of the lead sulphide, zinc sulphide and, in more recent times, iron sulphide, in the ore from the Sullivan mine. It is the author&apos;s intention to sketch briefly experiences encountered when the several types of flotation machines were introduced into the flow. Original Flotation Equipment and Ore The original flotation equipment consisted of four 18-cell lead-roughing machines, whose concentrate was cleaned in an 8-cell and recleaned in a similar 8-cell machine. Zinc concentration was carried out in four 18-cell machines. One 8-cell zinc re-treatment machine received the reground zinc middling. The 168 cells handled some 3000 tons of original daily feed. These cells were of the standard Minerals Separation type, equipped with 24-in. impellers. Spindles were driven through geared line shaft at a speed of 255 r.P.m. The impellers were the 45&apos; type, except for the last three cells of the zinc-rougher machines, which were of sub-aeration type, with full bottom shroud and half shroud for the top side of the impellers. Four-pound air was introduced. This recognition of the benefit of intensified flotation operation for "gleaning" was an important phase in the early operation and will be discussed further on in this paper. Sullivan ore during this period is described as a finely disseminated mixture of four sulphides, totaling go per cent made up of 13.6 per cent galena, 14.4 per cent mar-matite, I per cent pyrite and 61 per cent pyrrhotite. The specific gravity of the ore was approximately 4.4. There has been a gradual change in the ore content. The 10 per cent of nonsulphide constituents has increased to between IS and 20 per cent and all three sulphides have dropped somewhat. Added equipment and changing technique have made possible the treatment of 8000 tons per day of mill feed instead of the 3000 originally handled. Special Features of Flotation at Sullivan Sullivan flotation differs greatly from customary practice. The very heavy return of lead rougher middling and lead cleaner tailing, totaling at times 125 per cent of the new feed tonnage, has been the subject of much investigation. The practice is still employed. This operating condition is mentioned because it is the author&apos;s belief that the in-

Jan 1, 1947

By Chung Yu Wang, Guy C. Riddle

There are found in nature upward of II2 minerals containing antimony, but only a few of them, listed in Table I, can be considered as antimony ore-forming minerals. Stibnite (Sb2S3), antimony sulphide or glance, is the principal ore of antimony. The oxide ores—cervantite, kermesite, valentinite, sernarmontite—occur sparingly in nature. Dry methods are generally adopted for the extraction of the metal. Electrometal-lurgical methods have had much attention in America and Germany and have not yet found, on economic grounds, practical or extended application, except in a few places. During recent years, treatment of the poor grades of the ore, especially sulphide, by ore-dressing methods, has commanded some attention and been adopted at a few plants. The impure sulphide of antimony, resulting from the liquation process, is called lLneedle," "liquated" or "crude" anti- mony, and the refined metal itself is called antimony "regulus." The different processes for the treatment of antimony ores are shown in the diagram of Fig. I. Liquation of Crude Antimony The first step in the smelting of antimony is a simple one, the process of LLliquation," which produces crude or needle antimony. Ores containing $ 50 per cent Sb are used in the liquation process for the production of crude. The temperature required for liquation is between 550&apos; and 600°C. Ores to be liquated are broken to about walnut size. If the pieces are larger than this, the low heat used will not penetrate&apos; effectively, and if they are smaller the ore tends to pack too closely for adequate penetration. A packed charge also prevents the free escape of the fused sulphide. Intermittent Liquation in Crucibles The unique type of furnace that is in use in China for smelting is not highly efficient but is simple in construction and operation. The furnace is generally built

Jan 1, 1944

By G. Aertsen

THE CHAIRMAN (HENRY D. HIBBARD, Plainfield, N. J).-This excellent paper fills a gap in our Transactions and is most acceptable. In the early days another way of casting tire ingots, sometimes called cheeses, because of their shape, was to use a 22-gage sheet-iron cylindrical can for each ingot of the proper size to give the weight desired. These cans were bottom cast in groups, usually of four, each can being surrounded by a heavy iron ring somewhat taller than itself the top of the can being covered by a circular cast-iron plate having a small conical riser in the center, about 2 ½ in. in diameter at the bottom and 10 in. high. The top plate was clamped down to resist the upward pressure of the fluid metal and the space between the can and the outer ring, usually about 2 in. wide, was filled with dry sand. The riser permitted the escape of some of the air displaced by the steel and also increased the pressure in the steel when the mold was filled, and so diminished somewhat the tendency for gas holes to form. But little was known of segregation and pipe in those days, though the center of an ingot was understood to be the worst part, and it was separated in part in the disk which was punched out in perforating the ingot. This disk was about 8 in. in diameter and from ¾ to 2 in. thick, the variation in thickness being to rectify the weight. When an ingot was too -heavy, a thicker disk-was made by not driving the punch in quite as far as when the ingot-weight was right. Then the bloom was turned over and the punch driven in on the other side, which finished the punching and detached the disk. The tire bloom was then" becked" and later reheated and rolled as described in the paper. The forging was done with a 10-ton double-acting steam hammer. The steel generally speaking had lower carbon, from 0.45 to 0.55 per cent., than tires today have, and the tires were smaller, the most common size of ingot weighing 1,060 lb. each. BRADLEY STOUGHTON, New York, N. Y.-This paper was specially requested by the Iron and Steel Committee, because that Committee thought so important an industry should have its history written by one who was familiar with it, and before he had forgotten, or all the rest of the world had forgotten, its early stages. Mr. Aertsen has said that it is not an exhaustive research. I think it would not be any different if it were an exhaustive research, because our library made a careful attempt to find some written light on the history of these tires, but very little was found.

Jan 4, 1917

By Eugene W. Pearson

In a marked departure from conventional practice, the Power Plate process combines the two unit operations of leaching and electrowinning into a single continuous step for direct recovery of copper from sulfide concentrates, cement copper, copper scrap, and copper oxide slurries. Considering how environmental controls are pre- venting development of smelters for additional capacity or new construction, the Power Plate process provides a welcome pollution-free alternative to the smelting-electrorefining routes. And its low capital investment and low operating costs also recommend it to copper producers who are now using or investigating various hydrometallurgical methods.

Jan 12, 1974

By Oscar A. Glaeser

THE mine and plant of the United Verde Copper Co., located in Yav-api County, at Jerome, Arizona, have been described in various technical publications and, therefore, a brief outline of its essential features will suffice. The orebodies are of the schist replacement type, the main sulfide mass being a large lens-shaped body approximately 7 acres in cross-sectional area. In general, the mineralization is found along the iron-schist contact, with the orebodies extending about 1000 ft. along this contact and varying from a few feet to 250 ft. in width. The mine is located on a mountain slope and has openings into it on three horizons, as follows: 1. Raises and ore passes open vertically into the shovel pit. These extend down to the 1000-ft. level, where they terminate as loading bins in the Hopewell adit. The main fan intake air shaft, which also extends vertically down to the 1000-ft. level, has its surface opening on the hillside at the extreme north end of the shovel pit, and the main mine air return exhausts through a raise at the extreme south end of the shovel pit. 2. The 500-ft. level adit. 3. The Hopewell adit on the 1000-ft. level. The mine has a vertical downward extent of 2000 ft. below the 1000-ft. level. Underground operations are carried on through two main shafts, both of which are wholly underground. Shaft 5 is the ore-hoisting shaft. It has three compartments and extends from the 800 to the 3150-ft. level. Shaft 6 is the service shaft, and is used for lowering and hoisting men and material. It has a large cage compartment and three smaller compartments for counterweights, ladderway, and pipe and power lines. This shaft extends from the 500 to the 2550-ft. level, and is now being extended to the 3000-ft. level. An air shaft for the downward movement of fresh air from the fan has been provided, and extends to the 2550-ft. level at this time. Work has been started to extend this air shaft to the 3000-ft. level. A return airway for vitiated mine air, in the form of raises and short connecting drifts, extends downward to the 2100-ft. level. The underground mining operations extend from the 300 to the 3000-ft. level. There are about 45 stopes in active operation on the 17 levels

Jan 1, 1929

By Ralph Arnold

WITHIN the past three weeks the price of crude oil in the Mid-Continent field has been cut 50 .per cent. Similar action has been taken, or probably, shortly will be taken, in most of the other fields of this country. This cut is very drastic. Is it justified? Some of the causes advanced for the cut are: that it is in line with the industrial downward readjustment now taking place throughout the country, and that there is an over-production of oil, which can best be slowed down by lowering the price of crude. In part, these two reasons are logically advanced, but such underlying causes as the great production- in Mexico have not been prominently brought forward nor the fact that it would be advantageous at this time to fill up the empty storage of many of the consuming companies at a low &apos;price. In order to properly comprehend the situation let us compare for a moment the production- of oil in the United&apos; States and that in Mexico. Oil in the United States occurs in numerous often widely separated pools, varying in individual area, but each usually relatively extensive, compared to the area or gathering ground from which it is ultimately derived: It occurs in mod-erately porous reservoir rocks under such conditions of pressure, mobility and availability as to preclude its withdrawal except over a period of years-varying from 1 to at least 40 (average, say 15 years, depending on-the pool). One well will usually drain a limited area-2 to 10 acres-so that rate of withdrawal of oil under each owner&apos;s property is determined by the owner. - The important point to remember is that no matter how intensive the development, the oil reserves of -the United States cannot be materially exhausted&apos; for at least 20 years, and the total available supply, not including shale oil, will probably require 50 to 75 years, or longer, for its withdrawal, the yearly rate of production decreasing slowly from this time on.

Jan 1, 1921

Hermann A. Keller was born in Philadelphia, Pa., Mar. 23; 1860. He received a preliminary education at the gymnasium of Darmstadt, in Germany, and subsequently entered the University of Pennsylvania, graduating in 1881. After serving for a few months as an assistant instructor in mining and metallurgy at the University, he went to Leadville, Colo., where he became, at first, assayer and chemist, and in due time (1886) superintendent of the Arkansas Valley smelting works. In 1888, he was appointed metallurgist of the Globe works at Denver; in 1889, superintendent of the Philadelphia works at Pueblo; and in 1891, superintendent of the Parrot works at Butte, Mont., where he remained until 1896, in which year he was employed to reconstruct the smelting plant of the Mountain mines at Keswick, Cal. In 1897, he opened an office as consulting engineer in San Francisco; and from this time on enjoyed a wide practice both as metallurgical and as mining engineer. In the former capacity, Mr. Keller had already distinguished himself by the invention of the Keller mechanical roasting furnace employed at the Parrot works in Butte and the Germania works near Salt Lake. He was a frequent contributor to the technical journals, and he collaborated in the preparation of the chapter oil "Bessemerizing" in Dr. Peters&apos;s work, "Modern American Copper-Smelting." It is a curious coincidence that these two friends and collaborators died at almost the same time, Keller in Arizona and Peters in Massachusetts on the 16th and 17th of February respectively. The latter years of Mr. Keller&apos;s life were largely given to the examination and development of mines in all parts of the world-Alaska, Chile, Germany, etc. He finally established his professional headquarters in New York City. His death was caused by an underground locomotive accident at Clifton, Arizona, on Feb. 16, 1917.

Jan 5, 1917

By W. R. Hibbard

Although extensive pole figure studies have been reported by Brick1 and others2,3 showing preferred orientations in rolled and annealed 70-30 alpha brass, and by Hermann and Sachs4 in 70-30 alpha brass cups, comparatively little information has been published on textures in drawn 70-30 alpha brass tubes. Crampton5 reported a (110) , [111] texture in drawn brass tubes based on selected individual X ray photograms following the method of Norton and Hiller,6 but the extent and ease of developing this texture and the annealing texture have not been reported and pole figures of these structures were not found in the literature. The present investigation was further stimulated by the possible influence of textures on the mercury stress-corrosion cracking of brass tubes. According to Snoek,7 iron-nickel alloy sheet with a strong cubic texture was more resistant to corrosion than a randomly oriented sheet. Crampton5 reported that single crystals of 70-30 alpha brass drawn into tubes did not crack after 24 hr in a standard mer-curous nitrate test. Edmunds12 found that single crystals of 70-30 alpha brass did not crack in a standard mercurous nitrate test but did crack when subjected to ammoniacal vapor. Wasserman11 reported a similar failure of a single crystal of brass in ammonia. A tube with a strong texture might approach single crystal behavior and tend to resist mercury cracking if the differences in orientation between adjacent grains were sufficiently slight. Preparation OF Specimens Crampton&apos;s5 tubes drawn 28 pct reduction in area developed a slight degree of preferred orientation while in tubes drawn 48 pct it was more marked. Based on this information, in order to obtain specimens with markedly different degrees of preferred orientation but with the same final draw, two sets of 70-30 alpha brass tubes were made following the mill schedules shown in Table I. The final draw involved a large reduction in outside diameter, and a ratio of reduction in area to reduction in outside diameter similar to that reported by Crampton,5, 10 to cause mercury cracking of brass tubes. The brass contained 71.31 pct copper, 0.01 pct lead, 0.01 pct iron and zinc by difference. Determination of Pole Figures A 6-in. length from each tube was rotated rapidly by an electric stirrer in 25 pct nitric acid until the gauge was reduced to 0.003-0.005 in. The tubes were carefully cut longitudinally and the walls gently flattened. They were further etched in 25 pet nitric acid to 0.0015 in. thickness and (111) pole figures were determined using copper K-alpha radiation in the manner described by Brick.&apos; The three general types of pole figures

Jan 1, 1949

By W. R. Hibbard

Although extensive pole figure studies have been reported by Brick1 and others2,3 showing preferred orientations in rolled and annealed 70-30 alpha brass, and by Hermann and Sachs4 in 70-30 alpha brass cups, comparatively little information has been published on textures in drawn 70-30 alpha brass tubes. Crampton5 reported a (110) , [111] texture in drawn brass tubes based on selected individual X ray photograms following the method of Norton and Hiller,6 but the extent and ease of developing this texture and the annealing texture have not been reported and pole figures of these structures were not found in the literature. The present investigation was further stimulated by the possible influence of textures on the mercury stress-corrosion cracking of brass tubes. According to Snoek,7 iron-nickel alloy sheet with a strong cubic texture was more resistant to corrosion than a randomly oriented sheet. Crampton5 reported that single crystals of 70-30 alpha brass drawn into tubes did not crack after 24 hr in a standard mer-curous nitrate test. Edmunds12 found that single crystals of 70-30 alpha brass did not crack in a standard mercurous nitrate test but did crack when subjected to ammoniacal vapor. Wasserman11 reported a similar failure of a single crystal of brass in ammonia. A tube with a strong texture might approach single crystal behavior and tend to resist mercury cracking if the differences in orientation between adjacent grains were sufficiently slight. Preparation OF Specimens Crampton&apos;s5 tubes drawn 28 pct reduction in area developed a slight degree of preferred orientation while in tubes drawn 48 pct it was more marked. Based on this information, in order to obtain specimens with markedly different degrees of preferred orientation but with the same final draw, two sets of 70-30 alpha brass tubes were made following the mill schedules shown in Table I. The final draw involved a large reduction in outside diameter, and a ratio of reduction in area to reduction in outside diameter similar to that reported by Crampton,5, 10 to cause mercury cracking of brass tubes. The brass contained 71.31 pct copper, 0.01 pct lead, 0.01 pct iron and zinc by difference. Determination of Pole Figures A 6-in. length from each tube was rotated rapidly by an electric stirrer in 25 pct nitric acid until the gauge was reduced to 0.003-0.005 in. The tubes were carefully cut longitudinally and the walls gently flattened. They were further etched in 25 pet nitric acid to 0.0015 in. thickness and (111) pole figures were determined using copper K-alpha radiation in the manner described by Brick.&apos; The three general types of pole figures

Jan 1, 1949

By Charles Baker

THE district of Catorce, San Luis Potosi, ranks among the first half-dozen silver-producing camps f Mexico. Mining has been more or less continuous there for 150 years. The large producing mines, Mapum, Mazapil, Palomas, Musquiz, Minas Virjas, now for the most part idle or abandoned, are grouped around the head of two valleys: one, which drains to the west, falls 2000 ft. (609.6 m.) in 4 1/2 mi. (7.2 km.) between the main Catorce and the western foot of the range and is characterized by steep canon walls; the other, east of the main town of Catorce from which it is separated by a high divide, heads in a number of steep tributaries above the basin in which lies the town of Potrero and flows nearly north for 10 mi. (16 km.) to the northern base of the range; below Potrero, it has for the most part gently sloping valley walls and a broad valley floor. The Sierra Catorce is the highest range of the region, the summits reaching altitudes of over 11,000 ft. With the exception of a few cliffs of the erosion escarpment of the lower limestone and the canon walls below the main town of Catorce, which is the highest town in Mexico, the gently rounded slopes are characteristic of a maturely eroded mountain range composed of folded sedimentary rocks. As is usual in desert regions, the heavy-bedded limestones are the most resistant to erosive forces. Folded into anticlines, they form the mountain ranges; and the intermontane valleys and basins, originally carved out of less resistant shales, clays, marls and sandstones, are covered by debris forming a mantle with surface gradually rising upwards toward the hills. In time, the debris buries, in deposits formed from their own ruins, all but the highest summits of the original mountains; some f this region has already nearly reached this stage. Most of the debris is firmly cemented by calcium carbonate (caliche). Solution f the limestone by meteoric waters afforded the supply of lime carbonate, which has been redeposited upon evaporation of the waters as a cement binding the boulders, pebbles, and finer detritus.

Jan 9, 1921

By Russell W. Mumford

LITHIA compounds have a place in pharmaceutical and other chemical industries and lithium minerals in glassmaking and ceramics. The metal, although rare, is used to a minor extent in alloys. During the nearly 50 years of production of lithium minerals in the United States, the silicate mineral spodumene has continued as the chief source of raw material for the manufacture of lithium salts. Although twice as much lepidolite, a lithium mica, has been mined, most of it has been used in the making of certain types of glass. Amblygonite has been an additional but minor source of lithium compounds. Triphylite, petalite, and zinnwaldite have also been utilized in the past. All of these commercial lithium minerals contain low percentages of lithium (Table 1). In 1938, a new raw material for manufacture of lithium compounds made its appearance. This product was crude dilithium sodium phosphate (Li2NaPO4), obtained as a by-product of manufacture of potash and soda chemicals from the brine of Searles Lake, California. Searles Lake, which is now dry, lies in the extreme northwest corner of San Bernardino County, California, and consists of a large mass of salts. It was estimated by the United States Geological Survey that the crystal body contained about 25 pct voids, which are filled with a saturated brine. The brine consists mainly of the sulphates, chlorides, carbonates, and borates of sodium and potassium. Several other acid radicals, including phosphates, are present. Lithium is present in the brine in an amount corresponding to 0.01 15 pct Li2O. In the course of evaporation of water from the brine, during the process of manufacture of sodium and potassium chemicals, dilithium sodium phosphate crystallizes out and is recovered by a flotation operation. The lithium concentrate is filtered and dried and is then ready for shipment. It analyzes from 19 to 21 pct Li2O and now constitutes the highest known grade of crude lithium material. It is and will continue to be a major source of lithium for manufacture of lithium products. The variation in lithia content of the commercial minerals as mined is due to replacement of the lithia by soda and potash, to partial alter-

Jan 1, 1949

By C. R. Boyd

WITHOUT attempting to do more than give a preliminary or skeleton report upon the geology and minerals of Southwestern Virginia at this time, I am led to hope that the great commercial importance of the sixty miles of cross-section here crudely treated of

Jan 1, 1877

By C. R. Boyd

Without attempting to do more than give a preliminary or skeleton report upon the geology and minerals of Southwestern Virginia at this time, I am led to hope that the great commercial importance of the sixty miles of cross-section here crudely treated of

By F. Rosenblum, P. Spira, F. Kitzinger

The possibility of increasing the efficiency of mineral processing plants by means of computer control has prompted metallurgists to examine their flowsheets more critically in order to develop practical control strategies. Control schemes in use to date generally rely on the automation of traditional and previously manually regulated control loops, e.g., the addition of reagents to maintain a constant assay or a constant pH. This approach is predicated on the availability of well established primary sensors such as pH electrodes or on-stream analyzers. Noranda Research Center, in cooperation with Noranda Group concentrators, has developed a froth level monitor and a pulp density meter adaptable to this type of control. Although some of the basic principles on which they are based have been known for some time, their incorporation into the design of reliable, direct-reading instruments which can be interfaced with a computer is believed to be novel.

Jan 1, 1980

By F. Rosenblum, P. Spira, F. Kitzinger

The possibility of increasing the efficiency of mineral processing plants by means of computer control has prompted metallurgists to examine their flowsheets more critically in order to develop practical control strategies. Control schemes in use to date generally rely on the automation of traditional and previously manually regulated control loops, e.g., the addition of reagents to maintain a constant assay or a constant pH. This approach is predicated on the availability of well established primary sensors such as pH electrodes or on-stream analyzers. Noranda Research Center, in cooperation with Noranda Group concentrators, has developed a froth level monitor and a pulp density meter adaptable to this type of control. Although some of the basic principles on which they are based have been known for some time, their incorporation into the design of reliable, direct-reading instruments which can be interfaced with a computer is believed to be novel.

Jan 4, 1979