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By J. F Haseman, J. E. Davenport

In the brown rock phosphate fields of Tennessee there are large deposits of phosphate matrix in which quartz is a major constituent of the gangue, and which cannot be beneficiated by the conventional washing process to yield a concentrate of suitable grade for use in the electric furnace. A huge tonnage of high-grade phosphate concentrate is produced in the Florida pebble district from low-grade siliceous feed. However, the characteristics of the low-grade siliceous Tennessee ore are such that flotation under the conditions employed in treating the Florida ore1,2 results in inadequate beneficiation. Investigation of Flotation Methods Flotation has been used in 'Tennessee primarily to produce a concentrate of premium grade by moderate enrichment of a marketable grade of phosphate. 3,4 A secondary use was to increase the recovery of phosphate from the fines (down to approximately 200 mesh) from medium-grade ores with a concomitant moderate increase in over all grade.3,5 It was considered desirable to investigate methods of improving the grade of washed products from low-grade siliceous Tennessee ore when these products were substantially below the requirements for use in the electric furnace. The present paper describes the results of a laboratory-scale study of factors that affect the beneficiation of such siliceous phosphate ore by flotation. The primary objective of the investigation was to ascertain the conditions under which the ore could be made to yield a concentrate suitable for reduction in the electric furnace. It was assumed that a phosphate concentrate for electrothermal reduction should contain at least 26 pct P2O5 and should have a weight ratio of silica to lime of not more than 0.80. Since a concentrate of moderate grade can be used in the electric furnace, and since, as Grissom4 has shown in detail, the ore will not bear the cost of expensive treatment, the study of collectors was limited to the relatively inexpensive fatty acids. A secondary objective was to determine whether a portion of the phosphate could be recovered in a grade suitable for treatment with phosphoric acid to produce 45 pct superphosphate. The ore consisted of variably cemented sand interlayered with clay. The sand layers were composed of grains of quartz, phosphate, and clay embedded in a matrix of clay and phosphate. Some clay was finely disseminated in the phosphate grains. Little massive high-grade phosphate was present. The material used in the flotation tests was a washed sand that had been separated from the ore in a bowl classifier at the TVA phosphate-washing plant. The percentage composition of the sand was as follows: P2O5, 20; SiO2, 40; CaO, 29; Fe2O3, 2; Al2O3, 3; F, 2.3. The particle-

Jan 1, 1948

By J. F. Haseman, J. E. Davenport

In the brown rock phosphate fields of Tennessee there are large deposits of phosphate matrix in which quartz is a major constituent of the gangue, and which cannot be beneficiated by the conventional washing process to yield a concentrate of suitable grade for use in the electric furnace. A huge tonnage of high-grade phosphate concentrate is produced in the Florida pebble district from low-grade siliceous feed. However, the characteristics of the low-grade siliceous Tennessee ore are such that flotation under the conditions employed in treating the Florida ore1,2 results in inadequate beneficiation. Investigation of Flotation Methods Flotation has been used in 'Tennessee primarily to produce a concentrate of premium grade by moderate enrichment of a marketable grade of phosphate. 3,4 A secondary use was to increase the recovery of phosphate from the fines (down to approximately 200 mesh) from medium-grade ores with a concomitant moderate increase in over all grade.3,5 It was considered desirable to investigate methods of improving the grade of washed products from low-grade siliceous Tennessee ore when these products were substantially below the requirements for use in the electric furnace. The present paper describes the results of a laboratory-scale study of factors that affect the beneficiation of such siliceous phosphate ore by flotation. The primary objective of the investigation was to ascertain the conditions under which the ore could be made to yield a concentrate suitable for reduction in the electric furnace. It was assumed that a phosphate concentrate for electrothermal reduction should contain at least 26 pct P2O5 and should have a weight ratio of silica to lime of not more than 0.80. Since a concentrate of moderate grade can be used in the electric furnace, and since, as Grissom4 has shown in detail, the ore will not bear the cost of expensive treatment, the study of collectors was limited to the relatively inexpensive fatty acids. A secondary objective was to determine whether a portion of the phosphate could be recovered in a grade suitable for treatment with phosphoric acid to produce 45 pct superphosphate. The ore consisted of variably cemented sand interlayered with clay. The sand layers were composed of grains of quartz, phosphate, and clay embedded in a matrix of clay and phosphate. Some clay was finely disseminated in the phosphate grains. Little massive high-grade phosphate was present. The material used in the flotation tests was a washed sand that had been separated from the ore in a bowl classifier at the TVA phosphate-washing plant. The percentage composition of the sand was as follows: P2O5, 20; SiO2, 40; CaO, 29; Fe2O3, 2; Al2O3, 3; F, 2.3. The particle-

Jan 1, 1948

By J. F. Haseman, J. E. Davenport

IN the brown rock phosphate fields of Tennessee there are large deposits of phosphate matrix in which quartz is a major constituent of the gangue, and which cannot be beneficiated by the conventional washing process to yield a concentrate of suitable grade for use in the electric furnace. A huge tonnage of high-grade phosphate concentrate is produced in the Florida pebble district from low-grade siliceous feed. However, the characteristics of the low-grade siliceous Tennessee ore are such that flotation under the conditions employed in treating the Florida ore1-2 results in inadequate beneficiation. INVESTIGATION OF FLOTATION METHODS Flotation has been used in Tennessee primarily to produce a concentrate of premium grade by moderate enrichment of a marketable grade of phosphate.3,4 A secondary use was to increase the recovery of phosphate from the fines (down to approximately 200 mesh) from medium-grade ores with a concomitant moderate increase in over-all grade.3,5 It was considered desirable to investigate methods of improving the grade of washed products from low-grade siliceous Tennessee ore when these products were substantially below the requirements for use in the electric furnace. The present paper describes the results of a laboratory-scale study of factors that affect the beneficiation of such siliceous phosphate ore by flotation. The primary objective of the investigation was to ascertain the conditions under which the ore could be made to yield a concentrate suitable for reduction in the electric furnace. It was assumed that a phosphate concentrate for electrothermal reduction should contain at least 26 pct P205 and should have a weight ratio of silica to lime of not more than 0.80. Since a concentrate of moderate grade can be used in the electric furnace, and since, as Grissom4 has shown in detail, the ore will not bear the cost of expensive treatment, the study of collectors was limited to the relatively inexpensive fatty acids. A secondary objective was to determine whether a portion of the phosphate could be recovered in a grade suitable for treatment with phosphoric acid to produce 45 pct superphosphate. The ore consisted of variably cemented sand interlayered with clay. The sand layers were composed of grains of quartz, phosphate, and clay embedded in a matrix of clay and phosphate. Some clay was finely disseminated in the phosphate grains. Little massive high-grade phosphate was present. The material used in the flotation tests was a washed sand that had been separated from the ore in a bowl classifier at the TVA phosphate-washing plant. The percentage composition of the sand was as follows: P205, 20; Si02, 40; CaO, 29; Fe203, 2; A1203, 3; F, 2.3. The particle-

Jan 1, 1947

By H. Y. Bassett

THIS paper attempts to show the various major operations used in nonferrous tube mills and does not necessarily represent the current practices at the two plants of the Wolverine Tube Div., of Calumet and Hecla Consolidated Copper Co. Inc. The operations now utilized in the production of tubing are quite varied and show the continual interest and advance metallurgically and mechanically to produce the best quality at the lowest cost. Naturally, any advance or lowering of cost is dependent on large production runs, which mean longer machine running time with maximum worker productivity. Standardization of tube sizes and elimination of unnecessary alloys are therefore very necessary, otherwise the development of costly new production machines would be of little benefit to the consumer as a whole. It should be noted that in such a short presentation it is impossible to describe in any detail the major operations. The short descriptions, it is hoped, will enable all who are interested to obtain a general idea of present, day practices in the copper tube producing field. Fig. I is a flowchart or sequence of operations showing the steps usually employed in tube production. RAW MATERIALS Quality starts with the raw material, and it is absolutely necessary that raw material be controlled and held to rigid specifications. Little need be said, however, regarding the raw materials for the casting of bars or billets, as the material always is purchased and checked against the various governing A.S.T.M. specifications. CASTING Casting, as is indicated in Fig. I, is quite varied and for convenience can be classified into two groups: The casting of solid billets, and the casting of hollow tubes, each of which is the base for subsequent operations. In the melting of the raw materials various types of furnaces are used, and the selection of any one is dependent on many factors outside the scope of this paper. However, as of general interest, a short description will be made of each. Induction Furnace (Low Frequency) This is popular in the melting of copper-base alloys, Fig. 2. In the induction furnace the principle of an inefficient transformer is used for the production of the necessary melting heat. The primary coil is.arranged around a transformer core made of high permeability silicon steel sheet. The secondary circuit is through a melting channel around the primary coil and opens into a hearth which contains the main charge. The heat is generated only in the melting channel, but because of the stirring action there is a constant flow of metal around the channel, which when mixing with the charge causes very rapid melting. The volatilization losses are very low as the furnace atmosphere above the charge is heated only by the heat from the melting metal. The stirring action, however, makes the use of any slag cover difficult. Such furnaces- must be started with molten

Jan 1, 1951

By B. D. Harden

FOR over fifty years the Spanish mine, 21 miles northeast of Nevada City, in Nevada County, California, has been one of the Bradley properties. Between 1883 and 1889 it was operated by the late Frederick W. Bradley and P. R. Bradley. The ore averaged only about $1 a ton at that time but the mine was credited with a production of about $900,000 in that period. The property was then closed down until 1922, when the Bradley interests reopened it and continued development for the next nine years, when a 50-ton mill was built. This began operation in April, 1931. This was a selective flotation plant, with the tailing treated on Wilfley tables, and operated on ore from the No. 4 and No. 16 tunnel stopes. This ran from $3 to $6 in gold, 2 to 3 oz. of silver, copper 0.6 per cent, lead 1.5,

Jan 1, 1935

By T. B. Nolan, R. N. Hunt

In terms of present metal prices, analysis of extant records of the Eureka district indicate past production of the magnitude of $200,000,000 in recovered silver, lead, and gold. Production to date has been thoroughly oxidized ore from massive replacements, fissures, manta, and pipe-like bodies-in large part from the massive Eldorado (middle Cambrian) and Hamburg (middle to late Cambrian) dolomite horizons. In the productive area of about 6 square miles carbonate rocks comprise 80 per cent of the section, which includes early Cambrian to late Ordovician horizons. Four-fifths or more of the district's production came from a block of Eldorado dolomite in Ruby Hill that was moved in over younger horizons on strands of a regional thrust fault. The lower 500 to 600 feet of the block was much fractured and brecciated and became host to the ores. The Ruby Hill fault, a late normal fault, limits the Ruby Hill block of dolomite on the north for its entire length of nearly a mile and terminates the downward extent and exploitation of its ores. It introduced an hiatus of some 40 years in the district's development. Drilling of the past 20-odd years, followed by underground work, has now exposed the down-faulted extension of the Ruby Hill block of Eldorado dolomite and its contained ores, dropped 1800 feet in a dip shift to depths beneath the surface of 2000 to 3000 feet. The ores at depth are the sulfide counterpart of those mined in early years on the footwall side of the fault. The ores of the district may be attributed to the same source as the intrusive quartz diorite outcropping in a limited way south of Ruby Hill but found at depth elsewhere in drill holes. The long north-south axis of the district, its anticlinal structure, its major faults, and the distribution of the intrusives and, in their relation to the surface, the ores themselves, all suggest that erosion has not cut deep enough to expose the extent of intrusives and the ores consequent upon them.

Jan 1, 1968

By Charles R. Keyes

I. INTRODUCTORY. IN the dry regions of the globe many silver-deposits display certain remarkable features which at the same time are so totally unlike anything met with among ore-bodies elsewhere, that they have long presented exceptional difficulties not only to a satisfactory explanation of their origin, but to economical milling. A most striking peculiarity of these ores is the haloid character of many of their bonanzas. To this fact more than to any other may be ascribed most of the unusual uncertainties of mining-operations in desert countries-the marvelous local richness of the ores under certain conditions and the often abrupt transition, to utter poverty; the inadequacy of ordinary prospecting customs and the constant devising of new exploratory methods; the unaccountable losses in smelting and the continual change in treatment-practice. For these anomalies certain climatic conditions seem to be chiefly responsible.

Jul 1, 1911

By Samuel S. Wyer

THE installation of the gas-producer power-plant in America has been so unusual that all engineers have viewed it with in¬terest; a large majority, however, regard it with a lack of con-fidence and many with positive distrust. Despite the fact that European engineers have usually been less inclined to take the initiative along experimental lines than are Americans, they have, nevertheless, developed the gas-producer plant to a very high state of efficiency to which they were forced by the neces¬sity of economy in fuel-consumption. The gas-producer power-plant is so common in Europe that engineers as well as the general public regard it with the same degree of confidence that is now universally placed in steam¬plants. Gas-engines, both small and large, are in general use there, and central stations, aggregating several thousand horse¬power, are quite numerous. The fact that gas-producer power-plants have received so little attention in America may be attributed to five conditions:-1', ignorance and prejudice; 2, newness of work; 3, in-

Mar 1, 1905

By R. J. Wasilewski

Elastic constants have been determined on single crystals of maximum-melting-temperature NiAl compound (50.6 at. pct Al) at 25°C. Temperature variations of Young&apos;s modulus in the three principal crystal directions between -150°C and +800°C exhibit no anomalies similar to those repovted for the isomor-phous compounds brass and AuCd. It is concluded that the ordered CsCl structure is stable throughout the temperuture range investigated NICKEL aluminide is a congruently melting inter-metallic compound of CsCl (B2) structure, stable over a rather wide composition range (40 to 54 pct Al). It exhibits a lattice-parameter maximum of a. = 2.883A at 50.6 pct 1,&apos; at which composition the excess aluminum atoms are randomly distributed on the vacant nickel sites;&apos; at higher nickel compositions the excess nickel atoms replace substitutionally in aluminum sites, while at higher aluminum compositions vacancies are present in the nickel sublattice.&apos; For compositions close to NiAl no significant variation in order has been observed at temperatures up to 900°C.3 At high nickel (65 pct) compositions the existence of a martensitic structural transformation has been reported.4 NiAl is isomorphous with the ordered B-brass structure and with the high-temperature structures of AuCd and TiNi alloys close to equiatomic composition. All of these exhibit anomalous elastic behavior as the temperature of their diffusionless transformation to lower-symmetry structures is approached.&apos;-&apos; This behavior has been explained as indicative of low stability of the CsCl structure in these compounds.8 The present paper reports the experimental determination of the elastic compliances of single-crystal NiA1. EXPERIMENTAL METHOD 1) Specimens. Polycrystalline NiAl was prepared by induction melting of carbonyl nickel and high-purity aluminum and casting into solid copper molds in gettered helium atmosphere. The castings thus produced were canned in mild steel and impact-extruded at 1100 to 1200°C to 38-in.diam rod. After the steel sheath was pickled off, the rods were converted to single crystals by repeated high-frequency satisfactorily produced the desired orientations in the form of crystals 38 in. diam by 8-12 in. long. These were subsequently center less-ground to 7 to 8 mm diam by 12 cm long, electropolished (in 80 pct methanol, 20 pct HzSO, at 25°C and 15 v), and their axial orientation determined to 1 deg by the Laue back-reflection technique. The rod dimensions were determined to 0.01 mm and the densities determined from geometrical dimensions and the weight of every specimen. The average density was 5.905 0.005 g per cu cm. 2) Elastic Data Measurements. Room-temperature elastic compliances Sii were obtained by determining to 0.1 cps the fundameGtal resonance frequencies in the longitudinal and torsional vibration modes. At these frequencies the following relations hold: E =p(2lfif =pvl= Sn- 2r[Sll- S12- -£„«] g -P(2iftf = pvf = (s<« - fr(sii - sw - ±s»y where vl and vt are the velocities, and f andft the resonance frequencies of the longitudinal and torsional waves, respectively; 1 is the length of the specimen, and r is the direction cosine factor—{cos2 a • cos2 0 + cos2 a . cos2 y + cos2 0 ¦ cos2 Y}. Given three crystals of orientations (loo), (110), and (111)) i.e., F = 0, 14, and 13, we can obtain six equations for the three unknown compliances, thus providing adequate internal check. Mode mixing can readily occur in the relatively long wavelength resonance utilized here. This could cause a significant error in the torsion wave velocities calculated from the apparent resonance frequenies. Therefore, throughout the present work, a method, described previosl,&apos; was used to ensure that no significant mode mixing took place at resonance. The apparatus used was originally designed for the Elastomat, Magnaflux Corp.. Chicago, 111. determination of flexural vibration resonance frequencies. At room temperature torsional vibration mode could be established by suitable positioning of the thin wires from the input and pick-up transducers, and longitudinal mode by cementing thin magnetic foils at the ends of the specimen, and inducing the vibration by means of electromagnetic transducers. Only flex-ural resonance frequencies could, however, be obtained reproducibly over the temperature range of -150 to 800°C. These data are considerably less accurate, the error in the modulus being estimated at -0.6 pct from the specimen dimensions and the room-temperature Poisson&apos;s ratio. The reproducibility of the resonance frequencies in repeat runs was, however, to better than l in 15,000, hence the relative elastic modulus error less than 0.02 pct. Therefore, the temperature variation of the principal Young&apos;s moduli obtained by flexural resonance, relative to room-temperature values, is believed to be satisfactorily accurate.

Jan 1, 1967

By J. D. Cooner

SINCE my working life of 32 yr has been spent in and about the anthracite mines of the Hudson Coal Co., and the previous 4 yr in a college school of mines, I can write best about the safety program of a coal mining company. However, most of the suggestions, statements and practices mentioned herein are applicable to the members of a safety organization and program of almost any industry. It is realized that industrial officials actually interested in real safety work are fully acquainted with the requirements of a practical safety program. Therefore, for such persons there may be nothing new in this paper. The Hudson Coal Co. of Scranton, Pa., employs 7,000 men and produces 5,000,000 tons of anthracite annually at nine collieries scattered for a distance of 40 miles throughout the Lackawanna and Wyoming valleys of northeastern Pennsylvania. For the past 17 yr of the above mentioned 32, I have been engaged in safety work as a safety inspector, assistant mine superintendent, and at present the safety engineer. I have also had the privilege of visiting coal and metal mines in other parts of the United States and Canada. We Americans are the most careless people in the world. If any one doubts this statement why are 100,000 of our people killed accidentally annually? Or why during the almost four years we were engaged in World War II were more of our people killed accidentally at home than were killed in our army, navy, marine corps, coast guard, and other services during the same period? We are careless everywhere in our homes, on the highway, at work and at play. The purpose of this paper is to outline the requisites of a safety department and its program to stop the accidents occurring to employees at work in mines or industrial plants. The first requisite for a safety program is that the management must be convinced of the need and value of a real safety program. By "management" is meant each department head from the president down. Otherwise a good safety department with a good safety program will not succeed because it will not have the proper personnel and equipment to work with, its rules and regulations will not be enforced, and the members of the safety department will be belittled in the eyes of most of the workmen and supervisors. Next, a safety engineer and his inspectors should have worked in the production end of the mines as workmen and also in an official capacity. They are then in a position to have practical and first hand knowledge of the accident hazards of the industry, how to prevent them, and in general the problems of both the workmen and officials. Some men are appointed as safety inspectors when they are physically unable to carry on as mine officials. Such men may know the safety end of the industry and even be safety-conscious, but if they are physically unable to be mine officials they are also physically unable to be safety inspectors, as a safety inspector in our mines must walk from 2 to 4 miles on each daily inspection-and visit each working

Jan 1, 1948

By P. D. I. Honeyman

IN THE Southwestern copper region the event of greatest interest was the starting up of the new Hurley, N. Mex., smelter of the Chino Mines division of the Nevada Consolidated Copper Corp., which occurred in May. This plant, built as a wet-charge smelter for treating largely an all-concentrate charge, embodies in its de- sign the latest in materials-handling equipment. Noticeable around this smelter is the absence of tram cars and trackage for the handling of charge, dust and other materials. Concentrate from the mills is received in bottom- dump cars which are unloaded into a series of concentrate bins. Concentrate is then reclaimed through the bottom of the bins and discharged through a feeder on to a conveyor belt. Flux for the charge is fed on to the conveyor ahead of the concentrate charge. The belt-mixed charge is conveyed to the furnace building and discharged into the furnace charge bin ready for feeding to the reverberatory furnace. As in the McGill plant, the furnace is fed by vibrating conveyors, one down each side of the furnace.

Jan 1, 1940

By P. A. Beck, M. V. Nevitt

All binary and a number of ternary u alloys formed by first long period transition elements were examined and found to be ferromagnetic at low temperatures. The Curie temperatures for these alloys were measured. It was found that the correlation of the Curie temperature with composition may be described qualitatively in terms of the contributions of the various component atomic moments to the saturation moment of the alloy. Ternary Fe-Cr-Mo u phase alloys were also found to be ferromagnetic, with a Curie temperature dependent on the iron content and largely independent of the unit cell dimensions. Similar relations were found for the Co-Fe-Cr and the Co-Cr-Mo a alloys, where the Co + Fe content appears to determine the Curie temperature. RECENTLY, it has become increasingly evident that the u phase is some kind of an "electron compound." Several papers1-&apos; discussed the empirical correlation between the composition of the various u phases in alloys of the transition elements and the distribution of electrons in the partially filled d-levels of the component atoms. However, a picture of the latter can be formed only in a more or less intuitive manner, as in the Pauling theory," and the rigorous treatment of these problems is at present prevented by lack of adequate knowledge of the electronic structure of the transition elements in the metallic state. It appeared reasonable to expect that investigation of the magnetic properties, which are related to d-shell electrons, may help in clarifying the physical nature of the foregoing empirical correlation. Qualitative tests of samples of various compositions have shown that certain u alloys become ferromagnetic at liquid nitrogen temperature." The present investigation was undertaken to determine whether or not low temperature ferromagnetism is characteristic of a large variety of u alloys and to establish how the Curie temperatures of ferromagnetic u alloys are influenced by changes in composition and in lattice parameter. In the first portion of the in- vestigation, Curie temperatures were measured for a series of u alloys containing either vanadium or chromium with one or two of the elements manganese, iron, cobalt, or nickel. In the second phase of the work, data were obtained on the Curie temperature of ternary (Fe, Co, Cr)a, (Fe, Cr, Mo)a, and (Co, Cr, MO)U alloys, as a function of composition and of interatomic spacing. Experimental Procedures High frequency vacuum melting in high purity re-crystallized alumina crucibles was employed in the preparation of the u alloys, using metals whose lot analyses are given in Table I. Approximately two-thirds of the alloys were analyzed chemically, the analytical results being in acceptable agreement with the intended compositions, as seen in Tables I1 and 111. The specimens were annealed for homogen-ization in an atmosphere of purified 92 pct He and 8 pct H, mixture or in evacuated and sealed Vycor tubes at temperatures from 780" to 1300°C, depending on the temperature range of the u phase field for the particular composition. The annealing temwerature and time for kach alloy are given in Tables 11 and 111. Curie points were determined by an induction method; the change with temperature of mutual inductance between two coils was observed when the i-u specimen was inserted into the flux path between the coils. The transformer consisted of three coils each containing 1200 turns of 0.0050 in. diam enameled copper wire. The center coil, P in Fig. 1, served as the primary coil, and was linked magnetically with each of the two secondary coils, S, and S? which were connected in series opposition. The primary coil, P, was energized by an audio-oscillator, set at a frequency of 500 cycles per sec, while the signal from the transformer secondary circuit,

Jan 1, 1956

By J. D. Clark

ONE&apos;S lifetime does not usually offer the oppor-tunity to observe and be a part of the commercial development of an entirely new material. Petalite has been known for over a century, but at first it was presumed rare. However, it was found to be so abundant that it has become a new and extremely useful industrial tool. Mineralogy Petalite is a lithium-aluminum-silicate mineral. It was from petalite that J. A. Arfwedson first discovered lithium in 1818. The theoretical composition of Li2O • A12O8 • 8SiO2 is equivalent to 78.4 pct silica, 16.7 pct alumina and 4.9 pct lithia. The mineral in its known commercial occurrence is colorless to white and rarely gray, transluscent and rarely transparent. It is brittle, has a vitreous luster, and occurs usually in cleavable masses (fig. 1). It is classified as a di-silicate, but is similar in composition to the metasilicate spodumene, Li2O.-Al2O8-4SiO2 (table I). However, their specific gravities are quite different (3.13 to 3.20 for spodumene, 2.39 to 2.46 for petalite). Occurrence Petalite specimens have been reported in many localities; however, there are only three known countries having occurrences of commercial significance. The classic locality is Sweden&apos;s Utö Mines, which, although reportedly still containing considerable petalite, have been closed and are not expected to be operated in the near future. At Varuträsk, in northern Sweden, petalite carrying 4 pct lithium oxide is the main ore. In 1946 there was reported about 1000 tons on hand. However, the iron content was 0.60 pet too high for other than chemical use. Petalite has been reported in western Australia. At the Cottesloe mine, where small quantities have been reported available, it is produced in conjunction with feldspar. Petalite occurs in considerable quantity in the feldspar quarry at Londonderry, and it was estimated1 in 1943 that over 3700 tons might be available. The petalite in South West Africa was first discovered by F. J. Jooste, a lawyer by profession, in 1938. The following year he discovered a much larger deposit, and in 1948 another important discovery was made. These deposits occur near Kari-bib and constitute the only known commercial deposits. Their full extent has not yet been determined, but huge massive blocks of petalite of the purest grade, running into many thousands of tons, can be seen in all these deposits. To date, 60,000 tons of high grade, low iron containing ore has been blocked out in the major bodies. Mining for years to come will be confined to simple quarrying. Selection from accessory minerals has proved

Jan 1, 1951

By J. D. Clark

ONE&apos;S lifetime does not usually offer the oppor-tunity to observe and be a part of the commercial development of an entirely new material. Petalite has been known for over a century, but at first it was presumed rare. However, it was found to be so abundant that it has become a new and extremely useful industrial tool. Mineralogy Petalite is a lithium-aluminum-silicate mineral. It was from petalite that J. A. Arfwedson first discovered lithium in 1818. The theoretical composition of Li2O • A12O8 • 8SiO2 is equivalent to 78.4 pct silica, 16.7 pct alumina and 4.9 pct lithia. The mineral in its known commercial occurrence is colorless to white and rarely gray, transluscent and rarely transparent. It is brittle, has a vitreous luster, and occurs usually in cleavable masses (fig. 1). It is classified as a di-silicate, but is similar in composition to the metasilicate spodumene, Li2O.-Al2O8-4SiO2 (table I). However, their specific gravities are quite different (3.13 to 3.20 for spodumene, 2.39 to 2.46 for petalite). Occurrence Petalite specimens have been reported in many localities; however, there are only three known countries having occurrences of commercial significance. The classic locality is Sweden&apos;s Utö Mines, which, although reportedly still containing considerable petalite, have been closed and are not expected to be operated in the near future. At Varuträsk, in northern Sweden, petalite carrying 4 pct lithium oxide is the main ore. In 1946 there was reported about 1000 tons on hand. However, the iron content was 0.60 pet too high for other than chemical use. Petalite has been reported in western Australia. At the Cottesloe mine, where small quantities have been reported available, it is produced in conjunction with feldspar. Petalite occurs in considerable quantity in the feldspar quarry at Londonderry, and it was estimated1 in 1943 that over 3700 tons might be available. The petalite in South West Africa was first discovered by F. J. Jooste, a lawyer by profession, in 1938. The following year he discovered a much larger deposit, and in 1948 another important discovery was made. These deposits occur near Kari-bib and constitute the only known commercial deposits. Their full extent has not yet been determined, but huge massive blocks of petalite of the purest grade, running into many thousands of tons, can be seen in all these deposits. To date, 60,000 tons of high grade, low iron containing ore has been blocked out in the major bodies. Mining for years to come will be confined to simple quarrying. Selection from accessory minerals has proved

Jan 1, 1951

By John Skelton, E. J. Harvey

Seepage runs (stream flow measurements in a short time at many points in a basin,) and stream flow records are used to locate losing areas in stream basins in the Missouri Ozarks. The terrane consists of dolomitic limestone and sandstone of Ordovician and Cambrian age. Side-looking radar imagery emphasizes variations in topography and reveals two types of drainage basin topography, rough and smooth (deeply dissected and slightly dissected basins). In the deeply dissected basins (rough ones), main streams and tributaries generally gain. In the slightly dissected basins (smooth ones), the streams lose a part or all of their flow. Logan Creek, a tributary of Black River in the southeastern Ozarks, loses much of its storm runoff and all of its dry weather flow. East and West Forks of the Black River, north of Logan Creek, gain water through most of their main stems. On radar imagery Logan Creek basin has a light uniform tone and smooth texture in contrast to the darker tone and rough texture of the East and West Forks. The contrast is due to the difference in degree of dissection. Therefore, radar imagery may be useful in suggesting gaining or losing streams (or reaches of streams).

Jan 1, 1973

By Geoffrey Gilbert

A study of specimens shows that the key to both oxidation and enrichment at Duck-town is the behavior of pyrrhotite, which is in part dissolved and in part replaced by marcasite. Enrichment takes place chiefly at the expense of marcasite. THE material that forms the basis of this paper was collected in the spring of 1922, during a ten-day visit to Ducktown by the writer in the company of Prof. L. C. Graton. The time available for the work was far too short to allow anything like a thorough study of the deposits, and structural and stratigraphic questions, though they have an important bearing on the origin of the primary ores, were necessarily left untouched. Attention was focussed on the zone of oxidation and enrichment; and by devoting most of our time to the upper levels, open pits, and old workings, and examining the deeper levels only thoroughly enough to obtain a fair idea of the nature of the primary sulfide ore, we were able to make a reasonably complete survey of the various orebodies. With this field work for a background, the writer later made a study, in the Harvard laboratories, of the specimens obtained during the visit. As some of the facts brought to light by this study appear to be new, it seems worth while to set forth the outstanding features and a proposed explanation of them. Much of the Ducktown material that is obviously of greatest significance, as regards oxidation and enrichment, is soft and crumbly. Notwithstanding the desirability of gaining from such material all possible information as to its mineral composition and textural relationships, but little has heretofore been learned regarding it; but, by suitable preparation of this material in connection with the present study, it has been possible to examine it under the microscope without any disturbance of its texture and about as satisfactorily as if it were hard, fresh ore. The information thus gained has contributed measurably to the results set forth in this paper.

Jan 3, 1924

By M. J. Saarivirta

A high-purity copper-zirconium alloy system was imesti-gated. The zirconium content of the alloys studied varied from 0.003 to 0.23 pet. The solid solubility of zirconium in copper and some physical and mechanical properties of the alloys have been determined. By proper heat treatment, Cu-Zr alloy can develop a high electrical conductivity and resistance to softening at temperatures up to 500°C (930°F). This combination of desirable properties makes this alloy superior, to other com-merzcrrl copper base alloys for- use in the electrical conductor field. DEMAND for high conductivity copper alloys that retain their mechanical properties at elevated temperatures is growing rapidly. At present, copper-base alloys such as copper-silver and copper-chromium are used where these characteristics are required. As an example, commutators in electric motors are often made of these alloys. However, the life at elevated temperature of commutators made of copper-silver alloys is much shorter than the life of the same part made of copper-chromium and copper-zirconium alloys. Copper-silver alloys are subject to creep and subsequent failure when the service temperature exceeds several hundred degrees centigrade. Use of copper-chromium alloys has been known for the past several years, but little information is available in the literature concerning the properties of alloys in the copper-zirconium alloy system. Some work had been directed toward determining the solubility of zirconium in copper as well as determining properties of the Cu-Zr alloys. These investigations have revealed the superior high-temperature and electrical conductivity properties of copper-zirconium alloys over copper-chromium alloys. The United States Metals Refining Co., Carteret, N. J. has recently produced OFHC R copper-zirconium alloys for experimentation and evaluation. Because OFHC R copper contains no appreciable oxygen, no harmful oxides are formed and the full benefit of Zr as an alloying addition is obtained. This paper describes the results of this work. It contains room-temperature and some high-temperature data which have been obtained after various processing techniaues. Actual elevated temperature properties are being determined and will be reported in a separate paper. Preliminary results obtained at elevated temperatures suggest that the Cu-Zr alloy is superior to other high conductivity copper-base alloys. WORK PROCEDURE The high-purity oxygen free (OFHC R) copper-zirconium alloys with zirconium content ranging from 0.003 to 0.23 pct were cast into 4-by 4-in. wirebars by both continuous casting and conventional casting methods. Zr was added in the form of a master alloy containing 30 pct Zr and 70 pct Cu. Castings were analyzed for zirconium and nine bars containing various amounts of zirconium were selected for this study. The results of analyses and the sample numbers are given in Table I. Typical analyses of OFHC R copper are also shown. The following properties were studied: A) Solid solubility of zirconium in copper. B) Effect of zirconium on properties of copper. C) Effect of combined working and heat treating on properties of Cu-Zr alloys. The solid solubility of zirconium in copper was determined by metallographic techniques. Specimens for this phase of the investigation were hot rolled at 900oC (1650oF) to 0.250-in. rod. The rolled rods were cut into 2/3-in. long specimens and solution annealed at 950 °C (1740°F), 980°C (1795 °F). and 1020 °C (1870 °F) for 6 hr and quenched in water. In order to find the solid solubility limits at lower temperatures, the homogenized specimens were reheated at various temperatures for 1 hr and quenched. The physical and mechanical properties were determined on 0.250 in. rod, 0.132- and 0.081-in. wires. The wires were cold drawn from the 0.250-in. rods after they were solution annealed at 980 °C (1795°F) and quenched in cold water. Mechanical and electrical properties were determined on cold drawn and

Jan 1, 1961

By Arthur Notman

THE raw material men in all industries, and copper is no exception, are accustomed to think of them- selves as the whole show, and not without justice, for if there were no copper mines the world would have no copper. Nevertheless, many different things costing money and resulting in profits must be done to their product before it is ready for the world&apos;s use. In other words, they are making the frequent mistake of Chinking in terms of partial costs. Nobody really uses electrolytic copper. Possibly that is why nobody seems willing to pay anything for it. The manufacturers, being canny, do not buy any more electrolytic than they feel that they can work up and sell profitably, although they like to make the raw material men Chink that they will take all that can be produced. This is a state of affairs that has aroused my sympathy and curiosity, and has induced me to study the financial history of the copper wire and brass industry of the country or such portion of it as publishes any records readily available to the curious minded. About 60 per cent of the electrolytic copper used in this country is drawn into wire and consumed by the electrical industries. The remaining 40 per cent is

Jan 1, 1926

By L. D. Tracy

On Aug. 5 and 6,1918, and Mar. 26, 1919, the writer made an investigation of the mine-water neutralizing plant at the Calumet mine of the H. C. Frick Coke Co. The object of this plant is to develop a process of treating the water pumped from the mines so that it may be rendered suitable for use at the power plants of the mines and also at the coke ovens; at the same time, eventually producing a byproduct which will have enough commercial value to provide a sufficient revenue to place the plant on a self-sustaining basis. The Calumet mine is situated in Mt. Pleasant Township, Westmoreland County, Pa., on a branch of the southwest branch of the Pittsburgh Division of the Pennsylvania Railroad. It is about 6 mi. (9 km.) southeast of Greensburg, the county seat of Westmoreland County. The coal, which is the Pittsburg or Connellsville seam, lies at a depth of about 200 ft. (60 m.) and is brought to the surface by means of a shaft. The output of the mine averages 200,000 tons of coal annually, all of which is made into coke, either at the ovens, at the mine, or at byproduct ovens. There are 260 coke ovens at the mine. The continued development of the coal fields of Pennsylvania and the increased use of electric power in the operation of the mines has brought the problem of an increased water supply for the plants to the attention of the coal operators. This is especially true in the Connellsville coking region, where large quantities of water are used in quenching the coke at the ovens. Many of the streams receive the drainage from the mines; and as this water is highly acid and contains sulfur in various forms, some method of treatment is necessary to render it suitable for use at the plants of the mines. With this end in view, the H. C. Prick Coke Co., about 4 years ago, installed at its Calumet mine a plant for experimental purposes, in an endeavor to develop a process that would provide a maximum amount of suitable water at a minimum cost. From a purely technical point of view, the result of these experiments has been encouraging. The company is now endeavoring to place the plant on a commercial basis.

Jan 1, 1922

By Robert C. Stephenson

Large deposits of titaniferous magnetite occur associated with anorthosite and gabbro in the Lake Sanford area, Essex County, New York. The ore, gabbro, and anortbosite show consanguineous relations. Anorthosite grades through gabbroic anortbosite into gabbro but is locally intruded by gabbro. Structural and petrographic evidence suggests that gabbro was differentiated from anorthosite by filter pressing. Field and laboratory study of ore occurrences indicates that an ore residuum concentrated in the gabbro through magmatic segregation&apos; by normal crystallization and flowage. Ore occurs as bands which grade into gabbro, and also as large masses which replace anorthosite. Ore in the gabbro is finer grained and richer in ilmenite than ore in anorthosite. Introduction The Lake Sanford area lies in the heart of the Adirondack Mountains, in the southeast portion of the Santanoni Quadrangle (Fig. I), and there are four ore bodies in the area immediately surrounding Lake Sanford. The area lies 30 miles by highway from the railroad at North Creek. The deposits were discovered in 1826, and from then until 1858 iron was produced in small quantities. They were idle from 1858 to 1906. Between 1906 and 1913, extensive development work was carried on. These early attempts to produce iron ore were thwarted by transportation difficulties and by the titanium content of the ore, which, ironically, has been the element responsible for the present development. In 1941 the National Lead Co., Titanium Division, developed the Sanford Hill ore body and is now producing ilmenite and magnetite concentrates.18 The Defense Plant Corporation is extending the railroad to the property from North Creek and is building a sintering plant to treat the magnetite concentrate. Investigations OF the Deposits* Early studies of the Lake Sanford titaniferous magnetite deposits were made by Emmons in 1842,l2 by Kemp, in 1897-98&apos;0 and by Newland in 1908.22 Singewald26 and Osborne23 considered these deposits along with numerous other titaniferous magnetite deposits in 1913 and 1928, respectively. None of these reports entailed a detailed study of the ore occurrence in this area. The present study was begun in 1940. Seven months were spent in the field, and considerable time was spent in labora-

Jan 1, 1949