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By Larry G. Hayes

Development of the Bowers-Campbell zinc deposit, located in the Shenandoah Valley two miles northwest of Timberville, Va., was commenced in 1956 by the Tri-State Zinc Co. The occurrence of lead-zinc mineralization in this area has been known for many years, but the first serious exploration program was initiated in 1951 by the geologic staff of Tri-State Zinc Co. Ordovician dolomites. The deposit is approximately 700 ft long, 100 ft wide, and 475 ft thick. Its projected plan has an elongated ovoid shape, where- as the profile appears boot-shaped with the upper portion outcropping. Vertical cross sections show the lense as having been thrust-faulted along certain bedding planes. The movement along these bedding thrusts varies greatly and in one case, has offset the deposit so as to divide it into an upper and lower orebody. The breccia lense strikes parallel to the N 34" E strike of enclosing strata.

Jan 9, 1960

By Donald Ingvoldstad, Harold E. Lee

At Bunker Hill the original charge storage and preparation system was installed in 1917 to accommodate lead-silver gravity mill products. Only minor tonnages of wet fines such as vanner and flotation concentrates were received. Charge flux and diluent requirements were provided by an ample supply of siderite middlings and coarse limerock. While oversize material was crushed through Y4-in., the use of roll crushers in series resulted in a more or less granular, free-running charge of adequate porosity. Under such conditions, receipts could be proportioned directly from receiving bins, and suitable blending was obtained by passing the resultant layered composite through a small Stedman disintegrator. This elementary system served well for production and smelting of an extremely lead-rich sinter, in the low column blast furnace. Here sinter physical quality is less critical, and appreciable quantities of raw flux and oxidized ore may be charged directly to the blast furnace without seriously impairing furnace capacity.

Oct 1, 1956

By T. A. Wright

The sampling of waste materials containing copper, lead and tin has taken on a new significance within recent years, and is of increasing importance, on account of the entry of some of the copper refineries into this field and the tendency towards consolidations among the smelters. In the past, such material was sold on the basis either of a rough or grab sample or of casual inspection of the lot. The two principal sources of this waste are the processes of fabrication and the scrapping of automobile, electrical, plumbing, railroad and other equipment. In both cases much material can be sorted, salvaged and sold by classification and grade to be melted "as is" or with the addition of new or virgin metal. The increasing care in classifying exercised in the large manufacturing plants tends to increase the proportion of this usable scrap; on the contrary, the increasing use of more complex and special non-ferrous alloys tends to make sorting and grading more difficult in turnings, borings, clippings, etc., and consequently is making sampling of more moment. In tin, lead, aluminum, and zinc furnace by-products such as ashes, drosses, skimmings, etc., the trade practice of evaluation has not varied much. For instance, tin drosses, battery lead, and aluminum skimmings are ordinarily sold on what is called "metallic button" or "metallic recovery" or ('metallic yield." Copper and lead waste, however, is being sold more and more on actual copper and lead contents based on sample and chemical analyses. This paper will be devoted chiefly to the sampling of materials suitable only for refining and recovery of one or more of the metals present. The methods described are often based on the fact that but one metal, such as copper, is to be paid for. The loss of volatile constituents in melting does not appreciably affect the final copper result, as the "after-melting" assay can be readily calculated to the "before-melting" basis, but if smelter efficiency continues to increase, so that in time zinc, for instance, may be recovered on a paying basis, the method of sampling dirty brass, for example, will have to be changed also. As a matter of fact, we may expect that improved refining processes will make possible the partial recovery of lead and tin which are now lost.

Jan 1, 1928

By E. J. Duggan

THE mine and mill of the Climax Molybdenum Co. are at Climax, Colorado. Climax is on Fremont pass directly on the Continental Divide, at an elevation of 11,400 feet. DESCRIPTION OF ORE The Climax ore-bearing rock is essentially an altered and highly silicified granite, fully half of the gangue being quartz. Molybdenite is the only mineral of economic consequence and most of it is intimately associated with quartz in the form of fine veinlets and stringers. Other minerals are molybdite, pyrite, and chalcopyrite. While the molybdite may be readily extracted by hydrometallurgical methods, its content is too low to make such operations profitable. The pyrite content ranges from 2 to 5 per cent and that of the chalcopyrite from 0.03 to 0.05 per cent. There is not enough copper to make its recovery attractive and one of the metallurgical problems is the elimination of it and the pyrite from the product. On account of the dose association of the molybdenite with the quartz and the necessity of eliminating the pyrite and chalcopyrite from the final product, fine grinding followed by flotation is the only feasible method of concentration. HISTORY Construction of the first milling unit was started in 1917 and it was operated during parts of 1918 and 1919. In the postwar years, there was almost no demand for the product and production was at a standstill until 1924. The Company's research campaign to develop the use of the metal began to show results in that year. From 1926 to 1932 production ranged from 500 to 1200 tons per day. In 1931 an additional mill section was built and production was gradually increased until in 1934 and' 1935 about 3000 to 4000 tons was milled per day. This tonnage overtaxed the milling facilities, and as the demand for molybdenum was growing constantly, additional milling equipment was needed. In 1936 and 1937 another crushing plant, six additional mill sections and the necessary auxiliary plants were constructed and brought into operation. This plant was designed to mill 10,000 tons per day, but it was not long before 12,500 tons was being treated during periods of peak demands. During the summer of 1941, to meet defense needs, the mill has been required to treat 15,500 tons per day. This has been accomplished with little plant expansion and at no loss in recovery, as will be described later. To increase capacity still further, one more primary section is now being built. This will bring capacity to 18,000 tons per day, which can be stretched to 20,000 tons at a small sacrifice in recovery. CRUSHING There are two crushing plants: No. I, of 5000 tons capacity (flowsheet in Fig. 1) and No. 2, of 15,000 tons capacity (Fig. 2).

Jan 1, 1942

By E. J. Duggan

THE mine and mill of the Climax Molybdenum Co. are at Climax, Colorado. Climax is on Fremont pass directly on the Continental Divide, at an elevation of 11,400 feet. DESCRIPTION OF ORE The Climax ore-bearing rock is essentially an altered and highly silicified granite, fully half of the gangue being quartz. Molybdenite is the only mineral of economic consequence and most of it is intimately associated with quartz in the form of fine veinlets and stringers. Other minerals are molybdite, pyrite, and chalcopyrite. While the molybdite may be readily extracted by hydrometallurgical methods, its content is too low to make such operations profitable. The pyrite content ranges from 2 to 5 per cent and that of the chalcopyrite from 0.03 to 0.05 per cent. There is not enough copper to make its recovery attractive and one of the metallurgical problems is the elimination of it and the pyrite from the product. On account of the close association of the molybdenite with the quartz and the necessity of eliminating the pyrite and chalcopyrite from the final product, fine grinding followed by flotation is the only feasible method of concentration. HISTORY Construction of the first milling unit was started in 1017 and it was operated during parts of 1918 and 1910. In the postwar years, there was almost no demand for the product and production was at a standstill until 1024. The Company's research campaign to develop the use of the metal began to show results in that year. From 1926 to 1932 production ranged from 500 to 1200 tons per day. In 1931 an additional mill section was built and production was gradually increased until in 1934 and 1935 about 3000 to 4000 tons was milled per day. This tonnage overtaxed the milling facilities, and as the demand for molybdenum was growing constantly, additional milling equipment was needed. In 1936 and 1937 another crushing plant, six additional mill sections and the necessary auxiliary plants were constructed and brought into operation. This plant was designed to mill I0,000 tons per day, but it was not long before 12,500 tons was being treated during periods of peak demands. During the summer of 1941, to meet defense needs, the mill has been required to treat 15,500 tons per day. This has been accomplished with little plant expansion and at no loss in recovery, as will be described later. To increase capacity still further, one more primary section is now being built. This will bring capacity to 18,000 tons per day, which can be stretched to 20,000 tons at a small sacrifice in recovery. CRUSHING There are two crushing plants: No. I, of 5000 tons capacity (flowsheet in Fig. I) and No. 2, of 15,000 tons capacity (Fig. 2).

Jan 1, 1942

E. E. BACH, Ellsworth, Pa.-A personal interview with a man is one factor in reducing labor turnover; his working conditions is another, while the conditions under which lie brings up his family is still another; all have much to do with the reduction of labor turnover. A man well placed, happy in his work and happy in his home, will stay put and incidentally decrease the labor turnover in any industry. C. R. HOOK, Middletown, Ohio.-I am so deeply and sincerely interested in the employment problem, that it is very difficult for me, once I get started, to know when to, stop. In the first place, I want you to understand that I realize that what may be a possible accomplishment in the plant with which I happen to be connected, may not be possible of accomplishment in the same way in your plant. Our problems might be quite different; our location might be quite different, and therefore our employment problem would be very much affected by the community conditions. Two of our plants happen to be located in a town of about

Jan 4, 1918

DETAILED programs for the Regional Meeting to be held at Boston, on Aug. 29, 30, 31, will be available on arrival at the Copley-Plaza Hotel, where registration will begin at 10 a. m. on Wednesday morning. The opening session will be non-technical in character and will be called to order at 11 a. m. Hector J. Hughes, Dean of the Graduate School of Engineering of Harvard University, and others will speak, and in the afternoon visits are to be made both to Harvard and Massachusetts Institute of Technology. Wednesday afternoon will be devoted mainly to the opening of an exhibit of geophysical instruments and methods at Tech. On the way back to the hotel a tour of Old Boston will permit the visitors to see many of the historic points of the city. A special meeting of the Board of Directors has been called to be preceded by the usual subscription dinner at the University Club at 6.30 p. m. Owing to limita-tions of room, invitations have been sent only to the usual list designated in the By-Laws, the Directors, past officers, chairmen of sections and of technical committees. The dinner and a brief business session of the Board will be followed by a technical session at the Copley-Plaza at 8 p. m. and this in turn by a buffet supper at the University Club at 10 p. m. Sep-arate entertainment is being provided for the women for the evening. Thursday will be devoted to technical sessions both in the morning and afternoon. If necessary to permit adequate discussion, an overflow session will be ar-ranged in the afternoon. In the evening, a subscription dinner will be given at the Copley-Plaza at which past-presidents John Hays Hammond and Robert H. Rich-ards will be guests of honor. It is promised that speeches will be short and dancing will be arranged for following. The women will be offered a luncheon and bridge party at one of the country clubs on Thursday, and Friday will be devoted to an all-day excursion for both men and, women. Members are asked to register promptly on arrival so that the various committees will know as early as possible for how many to provide. The major part of the technical program has been provided by the Committee on Geophysical Methods of Prospecting, and a series of papers of great value and interest in this new field has been brought together. Abstracts of substantially all of them are printed in this issue of MINING AND METALLURGY and the papers themselves will, with one or two exceptions, be avail-able in printed form at the meeting and later in the regular publications of the Institute.

Jan 9, 1928

By Guillermo V. Borquez

INTRODUCTION The basic factors that influence the selection of a mining method are the geometry of the ore deposit and the strength of the ore and surrounding rock. These factors usually determine the ore recovery and dilution of the deposit under analysis. For the purpose of this paper we will consider a generalized case of a block caving system and the sequence of the analysis needed to optimize ore recovery. The analysis presented is intended for the preliminary planning stage. The critical factors involved in the caving process are identified and the methods for evaluating the suitability of a particular deposit to the caving method of mining are outlined. The mine design is approached by identifying the two most important, dimensions: the size and spacing of the draw points. These dimensions are estimated from a simulation model which takes into account the ore recovery, the incremental value of the ore recovered, and the safety conditions of the operations. The model estimates the number of draw points required to extract efficiently and safely the broken ore from the caving area. This number is found in the range where the net incremental value of the ore is maximized and the probability of pillar failure does not exceed the limit which is acceptable for a safe mining operation. A mining method and its design are in a continuous process of optimization. This process is unique for each deposit because of the many parameters involved; therefore, a back analysis of the operations should be performed when possible. These analyses will indicate if the ore body responds in a predictable manner to the parameters contained in the geological, fragmentation, and draw point models. SELECTION OF ALTERNATIVE MINING METHODS The selection of a mining method best suited to define the minable ore reserves of a given deposit involves consideration of the geometry of the ore deposit and the strength of the ore and surrounding rock. These two factors are usually sufficient to determine the ore recovery and dilution of the deposit. After analyzing the geological reserves and defining the limits of the deposit, either an open pit or an underground mining method can be selected to define the minable ore reserves. The sequence of steps to follow in the selection of a mining method is presented in Figure I. From an economic point of view, the strip- ping ratio is the determining factor for the selection between an underground or open pit method. In turn, the stripping ratio is a function of the grade distribution and con- figuration of the ore body. In order to obtain a preliminary estimation of the stripping ratio, a parametric analysis could be applied, such as the approach used by Professor E. P. Pfleider (1972) and summarized in Table 1.

Jan 1, 1981

By E. J. Duggan

The mine and mill of the Climax Molybdenum Co. are at Climax, Colorado. Climax is on Fremont pass directly on the Continental Divide, at an elevation of 11,400 feet. Description of Ore The Climax ore-bearing rock is essentially an altered and highly silicified granite, fully half of the gangue being quartz. Molybdenite is the only mineral of economic consequence and most of it is intimately associated with quartz in the form of fine veinlets and stringers. Other minerals are molybdite, pyrite, and chal-copyrite. While the molybdite may be readily extracted by hydrometallurgical methods, its content is too low to make such operations profitable. The pyrite content ranges from 2 to 5 per cent and that of the chalcopyrite from 0.03 to 0.05 per cent. There is not enough copper to make its recovery attractive and one of the metallurgical problems is the elimination of it and the pyrite from the product. On account of the close association of the molybdenite with the quartz and the necessity of eliminating the pyrite and chalcopyrite from the final product, fine grinding followed by flotation is the only feasible method of concentration. History Construction of the first milling unit was started in 1917 and it was operated during parts of 1918 and 1919. In the postwar years, there was almost no demand for the product and production was at a standstill until 1924. The Company's research campaign to develop the use of the metal began to show results in that year. From r926 to 1932 production ranged from 500 to 1200 tons per day. In 1931 an additional mill section was built and production was gradually increased until in 1934 and 1935 about 3000 to 4000 tons was milled per day. This tonnage overtaxed the milling facilities, and as the demand for inolybdenum was growing constantly, additional milling equipment was needed. In 1936 and 1937 another crushing plant, six additional mill sections and the necessary auxiliary plants were constructed and brought into operation. This plant was designed to mill I0,000 tons per day, but it was not long before 12.500 tons was being treated during periods of peak demands. During the summer of 1941, to meet defense needs, the mill has been required to treat 15,500 tons per day. This has been accomplished with little plant expansion and at no loss in recovery, as will be described later. TO increase capacity still further, one more primary section is now being built. This will bring capacity to 18,000 tons per day, which can be stretched to 20,000 tons at a small sacrifice in recovery. Crushing There are two crushing plants: No. I, of 5000 tons capacity (flowsheet in Fig. I) and No. 2, of 15,000 tons capacity (Fig. a).

Jan 1, 1943

By E. J. Duggan

The mine and mill of the Climax Molybdenum Co. are at Climax, Colorado. Climax is on Fremont pass directly on the Continental Divide, at an elevation of 11,400 feet. Description of Ore The Climax ore-bearing rock is essentially an altered and highly silicified granite, fully half of the gangue being quartz. Molybdenite is the only mineral of economic consequence and most of it is intimately associated with quartz in the form of fine veinlets and stringers. Other minerals are molybdite, pyrite, and chal-copyrite. While the molybdite may be readily extracted by hydrometallurgical methods, its content is too low to make such operations profitable. The pyrite content ranges from 2 to 5 per cent and that of the chalcopyrite from 0.03 to 0.05 per cent. There is not enough copper to make its recovery attractive and one of the metallurgical problems is the elimination of it and the pyrite from the product. On account of the close association of the molybdenite with the quartz and the necessity of eliminating the pyrite and chalcopyrite from the final product, fine grinding followed by flotation is the only feasible method of concentration. History Construction of the first milling unit was started in 1917 and it was operated during parts of 1918 and 1919. In the postwar years, there was almost no demand for the product and production was at a standstill until 1924. The Company's research campaign to develop the use of the metal began to show results in that year. From r926 to 1932 production ranged from 500 to 1200 tons per day. In 1931 an additional mill section was built and production was gradually increased until in 1934 and 1935 about 3000 to 4000 tons was milled per day. This tonnage overtaxed the milling facilities, and as the demand for inolybdenum was growing constantly, additional milling equipment was needed. In 1936 and 1937 another crushing plant, six additional mill sections and the necessary auxiliary plants were constructed and brought into operation. This plant was designed to mill I0,000 tons per day, but it was not long before 12.500 tons was being treated during periods of peak demands. During the summer of 1941, to meet defense needs, the mill has been required to treat 15,500 tons per day. This has been accomplished with little plant expansion and at no loss in recovery, as will be described later. TO increase capacity still further, one more primary section is now being built. This will bring capacity to 18,000 tons per day, which can be stretched to 20,000 tons at a small sacrifice in recovery. Crushing There are two crushing plants: No. I, of 5000 tons capacity (flowsheet in Fig. I) and No. 2, of 15,000 tons capacity (Fig. a).

Jan 1, 1943

By A. R. Troiano, R. F. Hehemann

The influence of partial decomposition to high temperature bainite on reaction kinetics at a lower temperature has been studied in two alloy steels. Reaction at the lower temperature is retarded by the prior treatment, and the extent of decomposition may be reduced. Interpretation of these results is based on a mechanism involving a limitation in the nucleation and growth of bainite plates. OF the major transformations in steel, the characteristics and general behavior of the bainite reaction are probably the least understood and appreciated. Limitations of space preclude a critical evaluation of the present status of the bainite transformation in this presentation; however, such a treatment will shortly appear elsewhere. Only the salient features pertinent to the present investigation will be introduced briefly here. Although the reaction curve for the formation of bainite is similar to that for a nucleation and growth process, other kinetic features are more in keeping with the martensitic mode of transformation. A definite temperature exists above which austenite will not transform to bainite.1-5 his temperature, which has been designated B., is determined by the composition of the austenite. Unlike other nucleation and growth processes, the amount of austenite transformed to .bainite is a function of reaction temperature. The extent of decomposition increases from 0 at H. to 100 pct at some lower temperature.' , This lower temperature will be designated B1 and appears to be relatively insensitive to austenite composition.% 5 The similarity in the effect of reaction temperature on the bainite and martensite transformations serves to emphasize the close connection between these two decomposition processes. Austenite decomposition in the bainite range proceeds without partition of the alloying elements.8-11 Partition of carbon has been proposed" primarily on the basis that partial transformation to bainite lowers M, and increases the amount of austenite retained at room temperature. Carbon enrichment resuslting from such partition has been employed to explain the influence of reaction temperature on the extent of decomposition.'" It should be noted, however, that no enrichment has been detected experimentally in high carbon steels.1,14,15 Lattice-parameter measurements of retained austenite in steels containing 0.3 to 0.4 pct C have indicated carbon enrichment, 3,10-18 although the split indicative of a high carbon martensite has not been reported. Carbon enrichment, if it does occur, must be highly localized around the bainite plates. Therefore, carbon enrichment does not account for the influence of temperature on the progress of the bainite reaction."' Thermal history is known to influence the martensite transformation through stabilization.20,21 No similar phenomenon in the bainite transformation has been reported. Materials and Procedure Two triple-alloy steels were chosen for this investigation. Their compositions were as given in Table I. These steels were chosen because the pearlite reaction did not interfere with the bainite reaction. Steel K was received in the cast condition and forged from 2 in. square bars to 1/2 x1 3/4 in. plates. The 4340 was received as 11/4 in. hot-rolled rounds. Both steels were homogenized in vacuum for one week at 2300°F in order to minimize segregation. A quenching dilatometer similar to that described by Flinn, Cook, and Fellows" was employed for the kinetic measurements. Dimensional changes were detected by a differential transformer coupled with a high speed recorder. The dilatometer was mounted so that it could be transferred to any one of three furnaces: a nitrogen-atmosphere austenitiz-ing furnace and two salt-bath furnaces for isothermal transformation. Dilatometer specimens were 1/32 x 1/4x 1/2 in. with a gage length of 1.4 in. All specimens were nickel plated in order to minimize decarburization during austenitizing. The austenitiz-ing conditions consisted of 10 min at the temperatures given above. Austenitizing temperatures were controlled to 210°F and transformation temperatures to ±3ºF. The precision of the dimensional measurements was estimated to be ± 5 x105 in. per in. Results and Discussion Isothermal Transformation: The characteristics of the isothermal bainite reaction will be described

Jan 1, 1955

By Edward Richardson

As DEVELOPED up to the end of the Great War, an airplane was essentially a mechanism of wood and fabric, joined and held together by metal fittings and, fastening. The engine and accessories, wire for wheel spokes and other tension members, and the occasional use of metal for interplane struts and the shafts of surface controls required some metal, but essentially it was a. wood and fabric structure. Attempts had been made to use metal for other purposes, such as struts in the drag-trussing of wing panels and for short interplane struts, and Junkers had produced an airplane of iron in 1915. By 1917 he was using steel tubes and duralu-min sheets, securing noteworthy advantages aerodynamically and from the standpoint of weigh PRINCIPLES OF AIRPLANE CONSTRUCTION The loaded weight of airplanes ranges from 500 lb. for the light plane of which we heard so much a short time ago, to about 40,000 lb. The great majority of planes range from 2000 to 10,000 lb. Of this weight from 15. to 30 per cent. is for the dividend-earning load. This percentage is so low that every per cent. saved on the rest of the weight increases the pay load from 3 to 6 per cent. The economic value of every pound that can be saved in designing or constructing a plane is sufficiently great to warrant spending money on refined methods of structural analysis, materials of superior strength properties for a given weight, methods of manufacture which may make possible the use of refinements in structural design, and the like, to an extent unknown heretofore.

Jan 1, 1928

INTRODUCTION TO COMPOSITING Individual drill hole samples may vary greatly in size, length, and weight. The sample interval may represent any convenient unit of hole length, and is usually related to the type of drill being used to obtain the sample. Depending upon the type of mineral deposit, it is desirable to group the individual sample values into composite weighted values representative of some particular zone or volume of the ore body. For example, values of several sequential samples of diamond drill core 3 m (10 ft) in length may be grouped together to represent a 12-m (40-ft) sample value over one bench height. On the other hand, several individual sample values taken across the thickness of a stratiform deposit or a vein zone may be weighted together to represent the value of the total thickness at the point of intersection. The choice of composite intervals is usually dictated by the type of mineral deposit and the mining method selected to produce the mineral. The techniques of weighting values from various lengths of samples into a composite value has already been discussed briefly in Chapter 3. Further examples of determining weighted composite values will be given in the following discussion on compositing for various types of mineral deposits. SAMPLE VALUE COMPOSITES AND TYPES OF MINERAL DEPOSITS There are many kinds of mineral deposits, each of which will require special treatment of sample data to obtain the best composite intervals for evaluating the deposit. It cannot always be known before hand what the optimum composite interval will be. Such composite intervals may properly be determined by geologic factors that are often unknown prior [ ]

Jan 1, 1980

By J. D. Forrester, Erich Sarapuu, T. C. Cheasley

John G. Tripp-—The concept of converting fuel deposits in-situ to useful forms of energy and chemical raw materials has important possibilities as a means of stretching out our fuel resources. The authors have worked in this field for years and in fact are among the pioneers in developing in-situ techniques based on underground electrocarbonization. Their efforts have aroused great interest in this subject and their contributions will undoubtedly gain even more significance as time goes on. The 4 to 5 billion tons of petroleum are proved reserves, that is, they are completely defined and could be produced economically with present methods of recovery. Reserves of coal, tar sands, and oil shale shown in Table I are not equivalent to proved reserves as the term is used in the oil industry. Incidentally, the 4.3 billion tons of oil shown in the tabulation is equal to about 28 billion barrels of oil and, therefore, presumably does not include reserves of natural gas liquids estimated at about 5 billion barrels. The relative comparisons would be more complete if these liquids together with the reserves of natural gas were included in the tabulation. It is stated in the report that production costs of the order of 10c per million Btu's could be expected from a large industrial electrocarbonization installation. The authors state that the operation probably could be operated successfully and at a profit. I would like to express my view that for the operation to be profitable extremely favorable conditions would have to prevail. For example, the coal seam would have to be thick and close to the surface in order to minimize drilling costs; the deposit should be dry and so situated that there would be little, if any, inflow of ground water; and the seam would have to be relatively gas tight to avoid excessive gas leakage. Notwithstanding the tremendous strides that have already been made, it is my hope that the authors plan to continue their development work. Experiments on underground electrolinking and initial carbonization at Gorgas, Ala., indicate a need for improved control over the operation. Subsequent examination of the underground system at Gorgas has shown that the gasification zone did not always follow a direct path between the electrodes. Undoubtedly the control of the in-situ thermal processes can be improved by further development work. With respect to the electrocarbonization of oil shale it may be well to point out that the cost of mining oil shale is estimated to be 50c to 60c per ton. These are low mining costs as compared to coal. Thus it would appear that there is less economic incentive for applying in-situ thermal methods to oil shale than there is for applying these methods to coal. As is the case with in-situ methods for petroleum, the product from the shale having the most economic interest would be liquid oil rather than gas. The authors point out that for distances up to several hundred miles it is cheaper to transmit energy as electricity than it is to ship energy as coal. This is a very important point, and with coal shipping rates rising periodically, it could prove increasingly important in the future. I believe that the authors deserve great credit for their outstanding work on underground electrocarbonization. The knowledge resulting from their efforts is of great value to our fuels technology. James L. Elder—The experimental work conducted by the authors on the electrolinking-carbonization of carbonaceous materials in situ is of great interest. The method is of great importance in the development

Jan 1, 1955

By Albert Sauveur

Floris Osmond, Honorary Member of the American Institute of Mining Engineers, born in Paris, March 10, 1849, died at Saint-Leu near that city, June 18, 1912. Taken suddenly ill with congestion of the brain while on his way from Paris to his home, he reached it in time to write on the typewriter his last wishes, one of which was that no biographical notice of his life should be published. Holding Osmond's memory sacred, as those who have known him will, is it our duty to conform with his desire ? Shall no mention be made in our Transactions of the beautiful life and great achievements of one of our most honored members? Is it not, on the contrary, incumbent upon us to disregard Osmond's instruction in this matter and to place on record our admiration for the nobility of his life, for his talents and for the importance of his work, that younger men may be inspired by it with a desire to be guided by worthy motives in the path of unselfish pursuits ? I cannot feel that in following this course I am disloyal to Osmond's memory and I am strengthened in this belief by the opinion of men of high honor whom I have consulted before undertaking the writing of this obituary notice. One of them, Professor Henry M. Howe, justifies my decision better than I could do it myself when he writes : " Let us look frankly at the matter. Here we have a most extraordinary life, a life of rare beauty and power of inspiration to the rest of mankind. In my opinion we should not deprive mankind of the stimulating influence of the example of M. Osmond simply because he shrank from publicity. Our young men see every day how greed and vulgar push succeed, and with what material rewards they are crowned. They see too rarely the more valuable crown which comes to self devotion. A permanent and public recognition of the beautiful qualities of Osmond's life cannot fail to appeal to the imagination of young men and to help to deflect them toward leading a noble rather than a sordid life. Therefore I feel that the Revue de Métallurgie and other bodies owe it to the public to give prominence to his career by means of biographical notices."

Jan 1, 1914

An old philosopher once said, "God grant me the serenity to accept those things I cannot change, the courage to change those which I can, and the wisdom to know the difference." It is time to take stock of how the human race has progressed-whether the trend now points toward greater evils or finer things. It is time that we stop fooling ourselves and put into effect a program of "real" conservation. By this it is not meant that we limit our concept to planting new forests, preserving wild life or preventing soil erosion, but rather taking stock of the expectancy of our resources, expanding human resourcefulness and effecting careful utilization and replacing of, or producing substitutes for, material resources. Man is still dependent upon energy and matter in certain usable forms, despite the fact that energy is inherent in all matter. Until such progress is made that the energy contained in all matter can be controlled, man must continue to utilize the common fuels. Advancement of research in atomic energy has cast a gleam of hope for future use of the tremendous energy contained in the atom. Use of atomic power for industrial pursuits would open up limitless resources of power. This energy, of course, is a two-edged sword, since man's progress and his means of destruction have always evolved parallel to each other. Thus it is estab-

Jan 1, 1950

By John M. Dale

Sulfur has attractive features as a base material for coating systems. It melts at 246ºF and can be applied as a liquid above this temperature. Its melting point is just above the environmental stress level, 180°F, for most coatings and yet it is low enough to be easy to work with in comparison to many hot melt systems. The viscosity characteristics are such that sulfur lends itself to spray applications. The oxidation rate of sulfur in the atmosphere is very low and sulfur has outstanding chemical resistance characteristics, particularly in acid environments. One of the greatest potentials for these coatings appears to be in the protection of other materials. At one time, most acid tanks and pits were constructed with acid- proof brick and sulfur based mortars. This type of construction is much more expensive than sulfur based spray coatings. An attractive sulfur based coating composed of sulfur, talc, milled glass fibers and dicyclopentadiene,1 developed by the Institute for the Bureau of Mines has excellent acid assistance and can be used to protect concrete from acid attack.

Jan 10, 1973

By Ingvar Janelid

13.7-1. General. Iron ore mining in western Europe is mostly carried out underground and only a minor part is mined by open pit methods. Actual open pits of importance in this connection are, among others, Svappavaara in Sweden, Mo i Rana and Sydvaranger in Norway, Corby Mines, Colsterworth, and Scunthorpe in Great Britain, and in western Africa, Nimba and Bong Range in Liberia, and Kedia d’Idjil in Mauritania. The open pit mines of western Africa are of great interest in this connection as they are designed according to European practice and have accepted the same techniques, methods, equipment, etc. as the European mines. In the surface mining of limestone, granite, shale, etc., the technical conditions generally are similar with those in the iron ore mines. Some information will therefore be submitted from typical quarries. The geological conditions at the various iron ore mines and limestone and other quarries vary extensively. In Scandinavia the iron ores occur exclusively in Precambrian rocks. The ore is hard, mostly magnetite or hematite, folded and steep dipping, and hence accessible to surface mining only in the shallow sections. In the mines referred to, open pit mining will last for another 10 to 30 years, whereupon mining has to be continued by underground methods. In France the Minette iron ore has previously been mined in open pits but, because of increased overburden, mining underground with room and pillar methods now dominates. The iron ores are Jurassic, flatlying strata of oolitic type, containing chamosite, siderite, and brown hematite. The British iron ores are mostly of the same origin and appearance as the French deposits, but despite heavy overburden they are preferably mined from the surface. In western Africa, the iron ores are of the same type as in Brazil, i.e., itabirites with high-grade portions resulting from silica-leaching and weathering. The ore contains mostly brown and blue hematite and magnetite. Nimba and Kedia d’Idjil are very similar regarding ore quality and geology. Bong Range is a deposit of low-grade ore which has to be concentrated.

Jan 1, 1968

By Sylvia Whitehead, G. M. Schwartz, J. N. Gundersen

Sylvia Whitehead (Petrologist, The Broken Hill Propriety Co. Ltd., Melbourne, Australia) — Examination of thin sections of jaspilites from numerous banded iron formations in Australia has revealed no evidence to suggest that silicate minerals have formed by reaction between magnetite and silica. Textural relations indicate that part of the magnetite crystallized very early in the history of the rocks, probably during diagenesis, and remained stable under conditions of metamorphism altering only to hematite or limonite under conditions of oxidation and weathering. Many jaspilites also contain later magnetite which is almost certainly of metamorphic origin. Wherever magnetite is present in quartz rich bands containing little or no silicate or carbonate, it is very fine grained and occurs generally as isolated, well formed octahedral crystals or as small crystalline aggregates. These show no evidence of corrosion or reaction with the quartz even in rocks where adjacent bands contain silicate minerals of moderate to high metamorphic grade. Concerning the silicate minerals, it is reasonably certain that bands of very fine grained stilpnomelane

Jan 1, 1963

By Sunder S. Saluja

Man had to learn to break rocks as early as the Stone Age, when they formed his main source of raw material. He started with chipping and over the years has reached a stage where he can employ atomic blasts and break, with one blast, millions of tons of rock. In between these two extremes there have been several stages of development including firesetting, use of the expansion of water on freezing in chipped cracks in cold climates, the use of low or deflagrating explosive black powder in about the 17th century, the use of high explosives with the discovery of nitroglycerine in 1846 by Sobrero, followed by the monumental discoveries of the great Swedish scientist Alfred Nobel in the latter half of the nineteenth century. The attempts to develop a stronger and still stronger explosive force were accompanied by investigations to understand the mechanism of failure of rock so that the force could be used to the best advantage; the latter, of course, have lagged far behind the former. This is mainly due to the fact that the chemical reactions involving high explosives as well as the processes resulting in the breakage of rock take place at too great a rate to be studied with the equipment that has been only recently available. Moreover, the development of stronger explosive forces was prompted by military necessities in the past and competition, at present, among highly developed nations. With the development, however, of the new techniques such as micro- second photography, microsecond flash radiography, extremely high-speed movies, strain gage assemblies, electronic chronographs and other improved electronic techniques, the investigator has in the last decade or

Jan 1, 1968