Search Documents

By R. E. Wagner

MISSOURI's famous lead area, in what is known as Southeast Missouri, is locally termed the "Lead Belt." These deposits are in the Bonne Terre dolomite of late Cambrian age which has a thickness of nearly 400 ft. Underlying this is the La Motte sandstone which varies in thickness from 0 to 400 ft. The La Motte beds are deposited directly on the ancient erosion surface of Pre-Cambrian porphyries and granites of probably Algonkian age. The buried igneous peaks extend up into the Bonne Terre in several areas within the Lead Belt. Overlying the Bonne Terre are shales and siliceous dolomites of late Cambrian and early Ordovician age. Within the Lead Belt these have a thickness of from 0 to 500 ft. The district is bounded in general on the southwest by the Sims Mountain fault, on the northwest and north by the Big River fault, and on the

Jan 1, 1947

Organization Place Date 1918 American Chemical Society Cleveland, 0. Sept. American Society of Sanitary Engineers Chicago, 111. Sept. National Petroleum Association Atlantic City, N. J. Sept. American Institute of. Mining Engineers Colorado Sept. 2-7 National Association of Stationary Engineers.... Cincinnati, 0. Sept. 9-13 Association Iron, Steel, Electrical, Engineers Baltimore, Md. Sept. 9-14 British Iron and Steel Institute, Autumn Meeting London, Eng. Sept. 12-13 National Exposition of Chemical Industries New York, N. Y. Sept. 23-28 American Electrochemical Society Princeton, N. J. Sept. 30¬Oct. 2

Jan 8, 1918

Organization Place Date 1918 American Chemical Society Cleveland, O. Sept. American Society of Sanitary Engineers Chicago, 111. Sept. National Petroleum Association Atlantic City, N. J. Sept. American Institute of Mining Engineers Colorado Sept. 2-7 National Association of Stationary Engineers.... Cincinnati, O. Sept. 9-13 Association Iron, Steel, Electrical, Engineers Baltimore, Md. Sept. 9-14 British Iron and Steel Institute, Autumn Meeting London, Eng. Sept. 12-13 National Exposition of Chemical Industries New York, N. Y. Sept. 23-28 American Electrochemical Society Princeton, N.J. Sept 30-Oct. 2

Jan 7, 1918

Organization Place Date 1918 American Chemical Society Cleveland, O. Sept. 9-12 American Society of Sanitary En Engineers Chicago, Ill. Sept. 3-5 National Petroleum Association Atlantic City, N.J. -Sept. American Institute of Mining Engineers Colorado Sept. 2-7 National Association of Stationary Engineers.... Cincinnati, O. Sept. 9-13 Association Iron, Steel, Electrical, Engineers Baltimore, Md. Sept. 9-14 British Iron and Steel Institute, Autumn Meeting London, Eng. Sept. 12-13 National Exposition of Chemical Industries New York, N. Y. Sept. 23-28 American Electrochemical Society Princeton, N. J. Sept. 30- Oct. 2

Jan 9, 1918

By Robert H. Richards

Jan 1, 1883

By T. J. Neil, C. B. Manula, O&apos

This paper describes a simulation model developed to analyze truck haulage problems as they are found in the open pit mining and construction industries. The model, using a computer, cycles trucks between their assigned loading and discharge points over a measured haulage route. Required parameters for model operation include the number and type of operating trucks and loading units, maximum truck acceleration and velocity, a realistic vehicular deceleration rate, and the ratio of ore to total material loaded at each origin. Equipment performance characteristics, system profile characteristics, and service time distributions are also included. Records are kept of all waiting times at the loading and discharge points and of any truck interference on the haulroads (moving queues), and a current journal of ore and waste production is maintained. In the pursuit of reduced truck haulage costs, management is confronted with three particular problems in truck and shovel mines. 1) What is the best assignment of trucks to shovels in the system? 2) If new equipment is needed, which new shovels and trucks will operate most effectively in the system? 3) Are any physical changes in the haulage layout warranted to increase productivity? For small open-pit mines the answers to these questions may be obvious. However, for larger operations complex problems are encountered with production and system configuration relationships usually being unknown. The mining simulator discussed here was developed to aid management in evaluating alternative pit-haulage schemes to reduce the risk involved in the selection and assignment of equipment in open-pit truck and shovel mining systems. There are two fundamental differences between this model and other truck haulage simulators: one is a difference of degree, the other a difference of kind. This model can simulate more complex systems than previous models by allowing transportation of material from multiple mine faces to multiple mine destinations and by being readily adaptable to a wide variety of mine layouts. The second difference that characterizes the model is the standard simulation of truck movements. This technique enables each vehicle to perform according to its mechanical capabilities and the physical profile of the haul road. Among the similarities between other truck haulage simulators and this model are the recognition of the probabilistic nature of certain service times in the system and the use of a digital computer to operate the model. The computer inherits the staggering computational and bookkeeping work loads and provides a means for compressing the time scale by allowing several days' operations to be simulated in a matter of minutes. DISTRIBUTION SYSTEM Basic aspects of the system analyzed here may be brought into focus by referring to Fig. 1, a schematic of the typical open pit truck-shovel operation programmed into the simulation model. A single stage materials handling scheme is employed to transport both ore and waste. Loaded trucks are directed to a crusher or stockpile if ore is being mined, while overburden is conveyed to a waste dump. Each shovel has a given number of assigned trucks and has two associated haulage routes, one for waste and one for ore. Up to 5 shovels and as many as 10 trucks per shovel (or a total of 50 trucks) can be operated in the model. Essentially the model, using a computer, cycles trucks between their assigned shovels and the proper discharge points over measured haulage routes. Required input parameters for model operation include the number of shifts to be simulated, operating time per shift, number and type of operating trucks and shovels, maximum truck acceleration and velocity, a realistic vehicular deceleration rate, and the ratio of ore to total material loaded at each shovel. Equipment performance characteristics, system profile characteristics, and service time distributions are

Jan 1, 1968

By Glenn S. Wyman

CHUQUICAMATA, located in northern Chile in the Province of Antofagasta, is on the western slope of the Andes at an elevation of 9500 ft. Because of its position on the eastern edge of the Atacama Desert, the climate is extremely arid with practically no precipitation, either rain or snow. All primary blasting in the open-pit mine at Chuquicamata is done by the churn drill, blasthole method. Since 1915, when the first tonnages of importance were removed from the open pit, there have been many changes in the blasting practice, but no clear-cut rules of method and procedure have been devised for application to the mine as a whole. One general fact stands out: both the ore and waste rock at Chuquicamata are difficult to break satisfactorily for the most efficient operation of power shovels. Numerous experiments have been made in an effort to improve the breakage and thereby increase the shovel efficiency. Holes of different diameter have been drilled, the length of toe and spacing of holes have been varied, and several types of explosives have been used. Early blasting was done by the tunnel method. The banks were high, generally 30 m, requiring the use of large charges of black powder, detonated by electric blasting caps. Large tonnages were broken at comparatively low cost, but the method left such a large proportion of oversize material for secondary blasting that satisfactory shovel operation was practically impossible. Railroad-type steam and electric shovels then in service proved unequal to the task of efficiently handling the large proportion of oversize material produced. The clean-up of high banks proved to be dangerous and expensive as large quantities of explosive were consumed in dressing these banks, and from time to time the shovels were damaged by rock slides. As early as 1923 the high benches were divided, and a standard height of 12 m was selected for the development of new benches. The recently acquired Bucyrus-Erie 550-B shovel, with its greater radius of operation compared to the Bucyrus-Erie 320-B formerly used for bench development, allowed the bench height to be increased to 16 m. Churn drill, blasthole shooting proved to be successful, and tunnel blasts were limited to certain locations where development existed or natural ground conditions made the method more attractive than the use of churn drill holes. Liquid oxygen explosive and black powder were used along with dynamite of various grades in blast-hole loading up to early 1937. Liquid oxygen and black powder were discontinued because they were more difficult to handle due to their sensitivity to fire or sparks in the extremely dry climate. At present ammonium nitrate dynamite is favored because of its superior handling qualities and its adaptability to the dry condition found in 90 pct of the mine. In wet holes, which are found only in the lowest bench of the pit and account for the remaining 10 pct of the ground to be broken, Nitramon in 8x24-in. cans, or ammonium nitrate dynamite packed in 8x24-in. paper cartridges, is being used. This latter explosive, which is protected by a special antiwetting agent that makes the cartridges resistant to water for about 24 hr, currently is considered the best available for the work and is preferred over Nitramon. Early churn drill hole shots detonated by electric blasting caps, one in each hole, gave trouble because of misfires caused by the improper balance of resistance in the electrical circuits. Primarily, it was of vital importance to effect an absolute balance of resistance in these circuits, the undertaking and completion of which invariably caused delays in the shooting schedule. Misfires resulting from the improper balance of electrical circuits, or from any other cause, were extremely hazardous, since holes had to be unloaded or fired by the insertion of another detonator. The advent of cordeau, later followed by primacord, corrected this particular difficulty and therefore reduced the possibility of missed holes. After much experimentation, the blasting practice evolved into single row, multihole shots, with the holes spaced 4.5 to 5 m center to center in a row 7.5 to 8 m back from the toe. Sucti shots were fired from either end by electric blasting caps attached to the main trunk lines of cordeau or primacord. The detonating speed of cordeau or primacord gave the practical effect of firing all holes instantaneously. Double row and multirow blasts, fired instantaneously with cordeau or primacord, proved to be unsatisfactory in the type of rock found at Chuquica-

Jan 1, 1953

By Glenn S. Wyman

CHUQUICAMATA, located in northern Chile in the Province of Antofagasta, is on the western slope of the Andes at an elevation of 9500 ft. Because of its position on the eastern edge of the Atacama Desert, the climate is extremely arid with practically no precipitation, either rain or snow. All primary blasting in the open-pit mine at Chuquicamata is done by the churn drill, blasthole method. Since 1915, when the first tonnages of importance were removed from the open pit, there have been many changes in the blasting practice, but no clear-cut rules of method and procedure have been devised for application to the mine as a whole. One general fact stands out: both the ore and waste rock at Chuquicamata are difficult to break satisfactorily for the most efficient operation of power shovels. Numerous experiments have been made in an effort to improve the breakage and thereby increase the shovel efficiency. Holes of different diameter have been drilled, the length of toe and spacing of holes have been varied, and several types of explosives have been used. Early blasting was done by the tunnel method. The banks were high, generally 30 m, requiring the use of large charges of black powder, detonated by electric blasting caps. Large tonnages were broken at comparatively low cost, but the method left such a large proportion of oversize material for secondary blasting that satisfactory shovel operation was practically impossible. Railroad-type steam and electric shovels then in service proved unequal to the task of efficiently handling the large proportion of oversize material produced. The clean-up of high banks proved to be dangerous and expensive as large quantities of explosive were consumed in dressing these banks, and from time to time the shovels were damaged by rock slides. As early as 1923 the high benches were divided, and a standard height of 12 m was selected for the development of new benches. The recently acquired Bucyrus-Erie 550-B shovel, with its greater radius of operation compared to the Bucyrus-Erie 320-B formerly used for bench development, allowed the bench height to be increased to 16 m. Churn drill, blasthole shooting proved to be successful, and tunnel blasts were limited to certain locations where development existed or natural ground conditions made the method more attractive than the use of churn drill holes. Liquid oxygen explosive and black powder were used along with dynamite of various grades in blast-hole loading up to early 1937. Liquid oxygen and black powder were discontinued because they were more difficult to handle due to their sensitivity to fire or sparks in the extremely dry climate. At present ammonium nitrate dynamite is favored because of its superior handling qualities and its adaptability to the dry condition found in 90 pct of the mine. In wet holes, which are found only in the lowest bench of the pit and account for the remaining 10 pct of the ground to be broken, Nitramon in 8x24-in. cans, or ammonium nitrate dynamite packed in 8x24-in. paper cartridges, is being used. This latter explosive, which is protected by a special antiwetting agent that makes the cartridges resistant to water for about 24 hr, currently is considered the best available for the work and is preferred over Nitramon. Early churn drill hole shots detonated by electric blasting caps, one in each hole, gave trouble because of misfires caused by the improper balance of resistance in the electrical circuits. Primarily, it was of vital importance to effect an absolute balance of resistance in these circuits, the undertaking and completion of which invariably caused delays in the shooting schedule. Misfires resulting from the improper balance of electrical circuits, or from any other cause, were extremely hazardous, since holes had to be unloaded or fired by the insertion of another detonator. The advent of cordeau, later followed by primacord, corrected this particular difficulty and therefore reduced the possibility of missed holes. After much experimentation, the blasting practice evolved into single row, multihole shots, with the holes spaced 4.5 to 5 m center to center in a row 7.5 to 8 m back from the toe. Sucti shots were fired from either end by electric blasting caps attached to the main trunk lines of cordeau or primacord. The detonating speed of cordeau or primacord gave the practical effect of firing all holes instantaneously. Double row and multirow blasts, fired instantaneously with cordeau or primacord, proved to be unsatisfactory in the type of rock found at Chuquica-

Jan 1, 1953

DeLamar, the largest silver mine in the United States outside the Coeur d’Alene District in Idaho, came into production in April 1977. It is a joint venture of Earth Resources Company and Canadian Superior Oil Co., with Earth Resources being the operator. The property is a consolidation of several old operations dating back to the 1800s; the better known of these are the DeLamar and Sommer-camp mines. DeLamar is situated in southwestern Idaho near the Oregon border in the Owyhee Mountains, 88 km (55 miles) southwest of Boise, Idaho. Heavy snowfalls and very low temperatures are common in the wintertime. Planned production was 77,500 kg (2.5, million oz) of silver and 5 13 kg (16,500 oz) of gold per year. Expected life at a milling rate of 1,540 mtpd (1,700 stpd) was 20 years, with an ore grade of 144 g Ag and 1.58 g Au per mt (4.2 oz Ag and 0.046 oz Au per st). Present production is at the rate of 1,900 mtpd (2,100 .stpd). The total cost to bring the mine into production was reported to be approximately $22 million, and the mill was reported to have cost about $10,000 per daily ton of design capacity. At DeLamar, naurnanite (Ag2Se) is the principal silver .mineral occurring with acanthite (Ag2S) disseminated through quartz veins and a silicified rhyolitic ash-flow tuff of Miocene age. The mineralization is relatively iron and manganese free and is readily amenable to cyanidation giving recoveries of 85% silver and 92% gold, which occurs free. When the plant was started, considerable difficulties were encountered in handling the wet sticky ore because of the high clay content. Following primary crushing, the ore is stockpiled. Due to the sticky nature of the ore, the vibrating feeders initially installed under the stockpile were replaced by a reclaim hopper and an apron feeder which was fed 24 hours a day by a front-end loader. The minus 200 mm (8 in.) ore is fed to a 5.5 m by 2.75 m (18 ft by 9 ft) autogenous mill followed by a 2.75 m by 4.57 m (9 ft by 15 ft) ball mill and cyclones. The mills are equipped with 750 kw and 520 kw (1,000 hp and 700 hp) motors. Both mills discharge to the same cyclone feed sump and a portion of the cyclone underflow is returned to the autogenous mill. Cyclone overflow, which is 20% plus 74 microns (200 mesh), is leached in four agitators in series for 72 hours (45 hours at the 2,100

Jan 1, 1981

By Henry Eyring, Jay W. Fredrickson

MANY theories have been advanced to explain the phenomena of elastic and plastic deformation. The object of this report is to present a mechanism for deformation, not radically differing from existing theories, but embodying certain principles and concepts of these theories. This proposed theory cannot be the complete answer to the age-old problem of deformation but it is believed that it will be a step further in the proper direction. The plastic properties of solids have recently been reviewed by Seitz and Read.1 Zener2 has reviewed and correlated the existing work on the anelasticity of metals. With these and othersa3,4,5 present theories of deformation are well-known. Of the many theories now extant, those receiving most attention have been proffered by Griffith,6 Becker,7 Smekal,8 Polanyi,9 Orowan10 and Taylor.11Although differing quite widely in details, these theories are based more or less on a common principle of flaws or dislocations that exist in the crystal lattice. The effect of stress concentrations surrounding a flaw or crack has been treated most recently by Fisher and Hollomon.12 The various authors differ as to the exact nature of the flaws or dislocations, the causes of their appearance, or the manner in which they initiate slip. All of the theories agree that the number of dislocations increases with progressing deformation. It is universally agreed that the shearing process is not a simple gliding of ideally perfect atomic lattice planes. It is also agreed that the "macroscopic process of gliding, or slip, is brought about by the spreading of a large number of locally initiated "glide steps" through the whole crystal. The occurrence of these "glide steps" is closely related to the presence of some kind of deviations from the "ideal" lattice structure, even in the undeformed crystal. These deviations are overcome "one at a time," in a manner of speaking, due to some stress-concentrating effect." PROPOSED THEORY Many solids show a continuously increasing deformation if subjected to a constant load of suitable magnitude in an extended range of temperature for a sufficient time. This deformation, or flow, may be said to be due, in general, to the tendency of the atoms or molecules in an elastically stressed body to readjust themselves in such a way that if the deformation were kept constant, a release of stress would take place. If rupture does not occur, this process of relaxation causes the rate of flow to obtain a final constant value, as shown in the curve (Fig I) obtained by Taylor and Quinney.14 In. this state the release of stress by relaxation exactly counterbalances the increase due to the elastic strain on the test pieces. .

Jan 1, 1948

For the year ending February, 1918 PRESIDENT PHILIP N. MOORE, ST. Louis, Mo. PAST 'PRESIDENTS WILLIAM L. SAUNDERS NEW YORK,. N. Y. L. D. RICKETTS NEW YORK, N. Y. FIRST VICE-PRESIDENT SIDNEY J. JENNINGS NEW YORK, N. Y. TREASURER GEORGE C. STONE, NEW YORK, N. Y. SECRETARY EMERITUS ROSSITER W. RAYMOND, _ NEW YORK, N. Y. SECRETARY BRADLEY STOUGHTON NEW YORK, N. Y.

Jan 11, 1917

By E. Swartzman, G. C. Behnke

THE free-swelling index, a numerical value for the free-swelling properties of coal, is being used to an increasing extent in specifying coal for burning in various types of equipment. The empirical standard ASTM method for determining this index uses, as a source of heat, a gas burner which requires frequent calibration because gas pressures are not constant, relatively small changes in gas pressure resulting in rather wide variations in the temperature attained in a crucible and thus in the resultant size of the coke button. Furthermore, the quartz crucibles specified for the test are not standard equipment, and relatively small variations in dimension and wall thickness from those specified often result in rather wide variations in the resultant size of the button. A method of analysis to be satisfactory not only for a scientific laboratory but also for use in industrial control laboratories under varying conditions should employ equipment which is rugged, easily calibrated, and readily replaceable by standard materials. Gas is not universally available as is electricity, and it is generally easier to calibrate and control a simple electric furnace or heater than a gas burner, the voltage not usually fluctuating to the degree that gas pressure does. Quartz crucibles are fragile, and the size specified in the test is not normally used for any other purpose in the laboratory. Although more expensive, the 15-cc platinum crucible, available in all laboratories where volatile matter in coal is determined by the standard ASTM method, is sturdy and standard and easily replaced. Because of these considerations, a modification of the standard method has been developed which employs a standard laboratory 350-w cone heater and standard 15-cc platinum crucibles. The results of numerous tests conducted with such apparatus show conclusively that the free-swelling index values as judged by the buttons produced with this equipment are exactly the same as those obtained by the standard gas burner method. Thus it is suggested that this electric cone heater method employing standard 15-cc platinum crucibles be considered as an alternative to the ASTM method, Designation: D720-46. The free-swelling index of coal is determined by means of the ASTM method, D720-46. This method, which is an adaptation of the crucible swelling test for coal of the British Standards Institution, consists of heating 1 g of coal in a special covered quartz crucible in a regulated gas flame for a period of not less than 2.5 min in which time the temperature of the inner surface of the bottom of the crucible reaches a temperature of 82025°C. The residual coke button is then compared in size and shape with a series of standard profiles shown in Fig. 1 and designated by number from 1 to 9 in increasing order of size. As the method is empirical, it requires rigid standardization of the apparatus and repeated calibration of the gas burner. The dimensions of the low form quartz crucible recommended must be rigidly adhered to as small variations in diameter or wall thickness result in variations in the rate of heat transmission and thus in the resultant size and shape of the coke button. The gas burner must be calibrated almost every time a new set of determina-

Jan 1, 1952

By Emile Gyss

THE speed of drilling rock has become an impor-tant factor in mining operations, while the place-ment of holes, kind, and quantity of explosive used. are equally important. These are a function of the rock hardness and toughness. Both of these fac-tors must be considered in compiling a table that will enable a mine operator or engineer to estimate drilling speeds in a given rock, knowing its hardness, and from its toughness, the dynamite required. The hardness of a homogeneous rock may be defined as its resistance to a penetrating medium, such as a drill-steel bit, and toughness, its resistance to with-stand rupture by the tearing apart of its minerals or parts of the same mineral. Rock hardness, as it has here been defined, is a function of the hardness of the component grains composing it, as well as grain size, shape, arrangement, and binding minerals. Rock tough-ness depends on substantially the same factors, but is measured by a totally different method. The units in which rock hardness is usually ex-pressed are either relative values or alphabetical desig-nations which carry no definite quantitative meaning. R. M. Raymond has suggested the measurement of com-posite rock hardness, based on mineral hardness as ob-tained by the Moh scale. This is done by multiplying the percentage of each mineral in a rock by the Moh mineral hardness and dividing by 100. These values convey a clearer conception in the minds of men in-terested in rock work -and are the basis of the units used in Tables 1 and 2. The hardness of rocks as it has been variously de-fined may be determined by different methods. The first method is one in which the proportionate average of the hardness of minerals making up a rock is the basis for classification. The other factors of grain size, shape, arrangement, and binding power are not considered.

Jan 6, 1927

By Horace C. Porter

Most of the bituminous and semibituminous coals of this country will coke, and all of them yield, on carbonizing, more or less marketable gas and byproducts. We need, however, a finer distinction as between various grades of coking coals and between coals of different degrees of byproduct and gas-making possibilities. There is required by the consumer or the coal producer a dependable test to show whether a certain coal or mixture of coals will make coke of good commercial value for blast-furnace, foundry, water-gas, or domestic use; whether any difficulties in operation may arise from properties inherent in the coal; and whether gas and byproducts may be expected of good quality and in profitable quantity. For determining coke quality, the most dependable test is an actual run in an oven, but for byproducts a coal may better and more accurately be tested in the laboratory or in a miniature oven or retort. As to what are the most desirable qualities in blast-furnace coke, the furnace men themselves are not in complete agreement. Some desire a hard, well-burned, "hot" coke, and others get better results from a llgreener," less hardened product that burns somewhat more readily in the furnace. But certain qualities of cell structure and porosity are commonly agreed upon as desirable. Information as to what kind of coke a certain coal may be made to yield is valuable in advance, whatever may be the consumer's ideas as to the particular kind most suitable to his needs. The question arises, Can we predict coking quality with any degree of satisfaction from a small-scale test? It used to be said that coals could not be tested satisfactorily on a laboratory scale either for the coke they would yield or the byproducts. On account of the mystery with which the coking quality of coal has been shrouded and the large determining influences that many believe are exerted by mass, pressure, and the passage of carbonaceous volatile matter through the coking material, it has seemed difficult to reproduce in the laboratory the exact conditions of commercial practice. Great progress, however, has been made in recent years in elucidating the process of carbonization and many of the factors that make for the formation of high-grade coke and the recovery of the best

Jan 1, 1920

By Robert P. Burrows

Discussion of the paper of R. P. Burrows, presented `at the New York meeting, February, 1916, and printed in Bulletin No. 107, November, 1915, pp. 2237 to 2245. EDWIN M. CHANCE, Wilkes-Barre, Pa.-I have been very much interested in the comprehensive and able discussion of the important topic of mine lighting just given by Mr. Burrows, especially as I know by common report Mr. Burrows' ability and his close touch with the development of the electric mine lamp, and its use in mines in connection with a portable electric lamp. It is regrettable that Mr: Burrows was not able to acquaint himself more closely with the cost of mine illumination, the actual cost, as produced by a portable light as carried by miners, whether a flame safety lamp or an open light. Had he been able to acquaint himself more closely with these data, he perhaps would not have arrived at some of the conclusions which he has presented to us in his paper. For example, we find Mr. Burrows advises that the open torch is by far the cheapest source of illumination, as produced by a portable light. I am very sure that is not the fact. The cost of miners' oil varies from about 25.c. a gallon to 90b. a gallon. The various States have passed laws from time to time rendering it mandatory that the miner shall use, whether he is willing or not, a so-called better or higher grade of burning oil. These laws have entailed the use of an oil so expensive that the use of the oil torch has become practically impossible, and the acetylene or carbide miners' lamp has largely taken its place. I have evolved a few figures showing the relative efficiency and usefulness of the open oil torch and the acetylene miners' lamp. For example, the acetylene miners' lamp consumes or burns 4 liters of acetylene per hour, or 0.14 cu. ft.; the oxygen consumed is 10 liters, or 0.35 cu. ft. per hour; air consumed, 50 liters, or 1.76 cu. ft. per hour; black-damp produced, 40 liters, or 1.41 cu. ft. per hour; air rendered extinctive to the acetylene flame, 112 liters, or 3.95 cu. ft. per hour; carbide con- sumed, 13 g. per hour. The open oil torch, on the other hand, burns about 20 g. of oil per hour; the oxygen consumed is 48 liters, or 1.69 cu. ft. per. hour; air consumed, 240 liters, or 8.47 cu. ft. per hour; blackdamp produced, 191 liters, or 6.73 cu. ft. per hour; air rendered extinctive to oil flame is 1,413 liters, or 49.92 cu. ft. per hour.

Jan 5, 1916

By B. S. Butler

A. D. Mutch (Falconbridge Nickel Mines Ltd., Ona-ping, Ontario)—This contribution is prompted by the fact that the writer has recently published an article'!' which has in part the same general conclusions as B. S. Butler's, namely, the form of the standard mineral paragenesis in hypogene deposits and the parallel transition from early high temperature acid to late low temperature alkaline or neutral solutions. Since the writer also recognizes that he is not fully qualified to discuss the formation of ore deposits as a chemical problem, this discussion will be limited to summarizing other observations and will suggest how Butler's Table II might be made more effective. The form of Butler's table is not unlike that widely used by European geologists'" in which the various minerals are related to the types of ore fluids—peg-matitic, pneumatolic and hydrothermal. Thus the decrease in temperature and change in pH of the ore-bearing fluids as recognized by Butler may also be paralleled by perhaps even greater changes in composition, density, and character. W. H. Newhouse's article:" on the sequence of hypo-gene ore mineral deposition is still one of the most complete analyses of the problem. In it he makes four major observations: 1) The order of O, S, As, Sb is the same as their increasing atomic weights and the same as their melting points. 2) The order of the metallic elements is the same as the Schuermann series. 3) The general mineralogical order is paralleled by decreasing heats of formation. 4) The order is generally in decreasing hardness, higher metal content, and higher specific gravity. The order of 0, S, As, and Bi is also the same as their relative abundance. Another point worth mentioning about this series is that As, Sb, Bi, when abundant, often substitute for sulfur in the early pyrite-like minerals, whereas later in the sequence these elements appear to act more like metals. This behavior may be a further indication of the acid character of the early fluids. C. J. Sullivan"' has attempted to show that the sequence of the metallic elements is related to the order of their melting points and has discussed the interrelationship of the different types of bonding of sulfo-phile and oxiphile elements in minerals. Items 3 and 4 might best be summarized by saying that bonding of the minerals in the series appears to be from ionic to covalent to metallic, the early minerals being light in weight, hard, and near-transparent; the intermediate minerals of intermediate weight, semi-transparent, and semi-conducting; and the end minerals heavy, soft, and malleable, opaque and good conductors. A rapid fall in temperature has been assumed to be the cause of ore deposition in many deposits. Perhaps this in turn is the effect of a rapid release of energy in the form of pressure release and change of volume, the first minerals to form being those with the highest heats of formation which absorb the released energy the most rapidly. Butler's Table II could probably be streamlined and made more effective if the listed compounds were assembled as mineralogical groups as in Dana. Thus the sulfides, arsenides, antimonides, bismuthides, and sulfa salts could be arranged in AX,, AX, A2X2, etc. groups. Since these are continuous substitution series, in which the fields of stability and the heats of formation of the various individual minerals can be fairly regarded as similar, this should be accomplished with no loss of clarity. The occurrence of the individual minerals and the precise sequence of the mineral deposition in any one deposit is probably more dependent on the concentration and relative abundance of the component elements. The assembling of geological data in tabular form such as Butler's Table II will continue to be a very effective method of studying geological phenomena. The importance of this particular problem has been pointed out by C. F. Park,% who concludes that understanding of mechanisms of the standard paragenesis and interrelated problems of mineral zoning in space and time is one of the major keys to classification of ore deposits.

Jan 1, 1959

By G. G. Biggar

A MATERIAL contribution to the success of our Armies in the field has been the portable pipe-line system. These are the words of Brigadier General R. F. Fowler, chief of the supply division of the War Department. The portable pipe line is one of the "secret weapons" the American forces employed in their swift advances through Tunisia, Sicily, and Italy. Constructed of a light steel so flexible it can be snaked around trees, it can be laid at a rate of ten to thirty miles daily right up to the fighting fronts. It has played a major role in supplying combat troops with products varying from petroleum to water.

Jan 1, 1943

By Stanley W. Sundeen

2.3-1. Introduction. The subject-"Preliminary Evaluation of Coal, Metallic and Industrial Mineral Deposits for Surface Mining'-seems at first glance to be definitive and not likely to be misinterpreted. Nevertheless, there are widely divergent opinions as to what is meant by "preliminary" and even as to what is meant by "evaluation." Some definition and qualification is necessary in order to provide a readily recognized basis for discussion. Evaluation evokes the idea of organizing extant data and judging its significance especially in the economic sense. It also may evoke the concept of the physical processes of acquiring the data that will characterize the deposit such as the geologic mapping, drilling, etc. Castle (5) considers that evaluation may embrace psychological aspects wherein human traits as, for example, over optimism might cause the evaluator to overlook obvious project weaknesses. In this chapter, evaluation is discussed primarily from the viewpoint of economically appraising a project from data available at a given point in time. The qualifying adjective "preliminary" is not easily defineable. In fact, some kind of evaluation proceeds, if not continuously, intermittently over the full span of exploration and early stages of development of a deposit. Bailly (1) comments: "Sometimes, before starting exploration in new areas or countries for a given commodity, we establish models to define what minimum combinations of re- coverable grade and tonnage are needed to give the minimum acceptable rate of return established by corporate policy. Then, before starting physical exploration in any target area we try and guesstimate optimistically what best 'geometry-grade-tonnage' can be expected. If these are acceptable in light of above minimum combinations, exploration is started as soon as funds are appropriated for the target area.''

Jan 1, 1968

By C. J. Huffman

MAGNESIUM alloy tubing and hollow shapes are finding increasingly wider usage in practical applications today. Square and rectangular tubing and hollow shapes find [ ] greater diversification in use than standard round tubing. Typical applications are roller conveyors, ladder rungs, spools, furniture, tank car loading tubes, athletic equipment, and business machine parts. In general, the amount of tubing and hollow shapes produced represent a minor portion of the magnesium extrusion business. The magnesium alloys in which tubing and hollow shapes are being produced together with typical properties are presented in Table I. MI A alloy is a low cost alloy possessing good welding and stress corrosion characteristics together with moderate formability and corrosion resistance. It is widely used for low strength applications. AZ31 B, alloy is used for moderately stressed applications and possesses good corrosion resistance and formability with fair welding characteristics. AZ61 A alloy has slightly higher physical properties than AZ31 B alloy together with good formabil- [ ] ity and corrosion resistance but slightly poorer welding characteristics. AZ80 A, ZK60 A-F, and ZK60 A-T5 alloys are used where the maximum in physical properties is required. AZ80 A alloy possesses fair corrosion resistance but with formability and welding characteristics inferior to' the other alloys. ZK60 A-F and ZK60 A-T5 alloys, while only recently offered for com¬mercial use, are finding favor due, to the higher permissible extrusion rates. These alloys possess greater toughness than AZ80 A alloy and, while of moderate formability at room temperature, are very formable at 300° to 500°F. The major portion of tubing and hollow shape production at the present time is in AZ3I B alloy. Magnesium alloy tubing is commercially available in the as-extruded condition or, in the case of ZK60 A-T5, in the as-extruded and aged

Jan 1, 1951

By Michael Tenenbaum

A REVIEW of process metallurgy of iron and steel during 1944 in many ways reflects the political and military developments of the year. Early in 1944 the tremendous wartime emergency expansion program was virtually completed and the industry directed its efforts to maintaining maximum production wherever needed. The vigorous production programs inaugurated to support the Normandy invasions did not persist as Allied armies approached Germany. General public apathy which developed with the prospect of an early ending of the European fighting was reflected in a general relaxation in armament programs. Toward the end of the year, as the prospect of a prolonged war became real, the entire program revived considerably and the need for steel became acute. Between January 1940 and October 1944 more than 12,400,000 tons had been added to the nation's steelmaking capacities, and over 12,700,000 tons added to the blast-furnace facilities. The total cost of this enormous program was more than two billion dollars. By July 1, 1944, total steelmaking capacity had exceeded 94,000,000 tons, with blast-furnace capacity rated at 68,400,000. Annual steel production, which was 88,900,000 tons in 1943, increased slightly during the past year and was estimated at 89,500,000. Production of pig iron also increased somewhat from 61,800,000 tons in 1943 to an estimated output of 62,500,000 in 1944. This review deals with a few of the numerous factors involved in maintaining the wartime iron and steel program at the level necessary to accomplish this enormous output.

Jan 1, 1945