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By L. E. Young

THE progress of mineral industry education will be limited to the period prior to World War II and will be considered as primarily a division of engineering education. Its relation to progress in the mineral industry and to the activities of the American Institute of Mining and Metallurgical Engineers will be emphasized. The history of mineral industry education has been recorded most completely and authoritatively by Dr. T. T. Read in his recent volume in the Seeley W. Mudd Series of American Institute of Mining and Metallurgical Engineers publications, entitled, The Development of Mineral Industry Education in the United States. Therefore, most of the historical details will be omitted from the discussion and an effort will be made to point out significant developments and to analyze trends. INTRODUCTORY STATEMENT AND SIGNIFICANT POINTS IN PROGRESS Among the significant points in the development of mineral industry education in the United States since 1871, the following deserve particular mention: I. The number of institutions offering undergraduate instruction has increased, particularly during the years 1890 to 1925; the number of students enrolled in undergraduate study has increased greatly, largely on account of the development of courses in petroleum engineering and increasing interest in metallurgy. 2. The selection of students to be admitted to undergraduate courses has been improved through the advisement, counseling, and guidance programs sponsored by the Engineers' Council for Professional Development. 3. Improved instruction methods and the training of the teaching staff have been promoted by the educational institutions themselves, but with the most effective stimulus and guidance of the American Society for Engineering Education and the Engineers' Council for Professional Development. 4. The accrediting of undergraduate curricula through the Engineers' Council for Professional Development, considering both qualitative and quantitative criteria, has done much to improve standards of instruction in engineering education for the mineral industry. 5. Undergraduate curricula and courses are revised currently to meet the needs of the American mineral industry. New curricula are developed currently to permit specialization in technical fields, particularly in the fourth year of undergraduate work. 6. There has been a strengthening of the fundamental courses in mathematics, physics, mechanics, and chemistry, and at the same time these courses are being adapted to the most recent developments in the industry, including metallurgy, geophysics, petroleum engineering, petroleum technology, ceramics and so forth. 7. There has been continuing improvement in plant and laboratory facilities of the educational institutions.

Jan 1, 1947

By A. C. Richardson

TROUGHS or launders are probably the oldest machines used for ore. concentration, and their development was suggested no doubt by the natural segregation and stratification of materials that take place in the bed of a stream. Trough washers for coal cleaning were first used in France and Germany about one hundred years ago and have been used in some form or other ever since. The early launders were intermittent or batch machines, and only comparatively recent developments have made them continuous. The chief disadvantage of early launder systems as applied to coal has been that when a satisfactorily clean: coal was produced the loss of good coal in the refuse was excessive. Other disadvantages have been the large quantities of water and the great lengths of troughs necessary to treat moderate tonnages of feed. One of the most widely used types of trough washers is the Rheolaveur, which has been developed in the last 25 years. Fig. 1 shows a more or less standard arrangement for fine coal. In this there are four troughs placed one below the other and sloped a few degrees from the horizontal. Rheo-boxes or draws are provided at intervals in the bottom of the trough for the removal of material deposited from the stream. These draws are spaced substantial distances apart. Some of the rheo-boxes are provided with upward currents of water to modify the horizontal current stratifica-tion. In operation it is the function of the first two troughs to remove all refuse material and permit only clean coal to be discharged over the end. In order to accomplish this, large quantities of coal are drawn through the rheo-boxes. The third and. fourth troughs are used to recover the coal discharged through the boxes of the first two troughs as a middling product and to discharge a clean refuse. The middlings are returned to the first launder and are usually about equal to 40 per cent of the raw coal being fed to the launder.

Jan 1, 1938

By Danforth R. Hale

Minerals for electronic and optical uses divide easily into two sections: (1) quartz and (2) minerals other than quartz. Quartz Quartz, having a great usefulness discovered by the radio communication enthusiasts about 1918, has become a subject of amazingly expanded interest in the U.S. over the past decade as a man-made mineral and a $3 million crystal-growing business. An economic summary of the commercial growing of quartz crystals has a place in a handbook directed to the mineral engineering industry when it is remembered that quartz crystals have long been an important commercial mineral, and that the raw material for “cultured quartz”- that is to say, quartz crystals grown through the ingenuity of man-is still natural quartz. Nearly all the natural crystals come from Brazil. The larger pieces which meet rigorous standards of quality are used for electronic and, to a lesser extent, optical components. Smaller pieces and fragments are used for vitreous silica. The need for high quality material in quantity led to U.S. government-sponsored research programs in the 1940s that have resulted in the factory growing of beautiful crystals of prescribed shape, size, and quality. The development of the cultured quartz crystal is typical of man's success in adapting the product of the mine to increasingly sophisticated uses. A remarkable achievement perhaps, but foreshadowed by experiments in 1908 by Giorgio Spezia (1908), an Italian geologist studying the relative effects of temperature and alkaline environment on the solubility of quartz. Modern radio equipment is most often controlled as to frequency by the presence in the circuit of a separately added "crystal"-the 1918 discovery responsible for the existence and growth of the quartz industry. The "crystal" is quartz, but this component is a carefully oriented and prepared slice from a crystal, but not a crystal as recognized by a rock hound or seen in a museum. How quartz operates to control frequencies is not a proper subject for a handbook on industrial minerals, and references should be consulted (Cady, 1964; Mason, 1964). Quartz belongs to that class of materials called dielectrics: those that do not conduct an electric current but permit electric fields to exist and act across them. Quartz shows the "piezo-electric effect," which means that when a quartz plate is mechanically deformed against its natural stiffness, one of its surfaces becomes negatively charged, the other, positively charged. When the plate is released from the stress quickly, the charges disappear as the plate regains its original shape, but because of mechanical momentum the plate deforms in the opposite direction (to a lesser amount) and the surfaces correspondingly become charged in the opposite direction. By coating the two surfaces thinly with metal and attaching flexible wires, these charges may be brought into an electronic circuit. If the surfaces are suddenly electrically charged by movement of current through the wires, the "converse piezoelectric effect" occurs -the plate deforms. Carry the thought further and it is realized that an alternating current flowing through the wires responds to the mechanical oscillation of the plate and is keyed to this oscillation. By controlling the thickness of the plate its mechanical vibration frequency can bevaried through a wide range. One type of quartz plate, called the AT-cut, having a precisely defined orientation with respect to the crystallographic axes of the crystal, vibrates on a microscopic scale much as a book would deform when placed flat on a table and the top cover moved parallel back and forth with the hand. There are at least 17 other orientations which have been studied, some of

Jan 1, 1975

By Ralph Sweetser

THE Round Table. on Qualities of Pig Iron, under the auspices of the Joint Committee on Qualities of Pig Iron, which is made up of members of the American Institute of Mining and Metallurgical Engineers, the American Foundrymen's Association, the American Society for Testing Materials, the Gray Iron Institute, and the. Malleable Iron Institute, was held at the Palmer House, Chicago, Ill., on Oct. 1, 1935. PROCEEDINGS R. H. SWEETSER,* New York, N. Y.-Seven years ago last June we had a round table on qualities of pig iron1, and at that time the foundrymen were out with their battle-axes for the blast-furnace men, and the open-hearth men were doing the same thing. After that meeting there was really a determined effort to find out about the qualities of pig iron, and great progress has been made all along the line. This afternoon we want to have a frank discussion of the qualities of pig iron, whether it is the molten pig iron going into the steel plant, or cold pigs going to the different foundries. We ought to hear something from Dr. J. T. MacKenzie, who is chairman of the Joint Committee, and from Mr. R. E. Kennedy, the tech-nical secretary. J. T. MACKENZIE, t Birmingham, Ala.-As a result of that meeting in Chicago, there was considerable agitation in the Foundrymen's Association to develop a set of tests for pig iron. One of the difficulties was that nobody talked the same language on shrinkage, draws, live iron, dead iron, and so forth, which were the main com-plaints against pig iron. By request of a committee of the Foundrymen's Association, the Bureau of Standards ran a comprehensive set of tests that did develop the prin-ciples governing the optical pyrometer as applied to cast iron and pig iron. Then the Bureau went on further with the fluidity test, and developed a method for the absolute changes of volume at the freezing point.

Jan 1, 1936

By F. H. Deily, L. W. Ledgerwood, J. E. Eckel

A new method for drilling the earth's formations has undergone extensive laboratory and shallow earth drilling experimentation. This method, pellet impact drilling, uses the energy of a large mass of high velocity steel spheres to break and crush the rock. Pellets are confined to a region close to the hole bottom. Motive fluid for transporting and recycling the pellets is supplied by a modified jet pump. The pellet drilling method has inherent advantages over conventional rotary drilling. First, only a nominal amount of weight need be applied to the hole bottom. Second, only a very small amount of torque is required for bit rotation. Third, means are available for continual replacement of the cutter medium. The experimental laboratory work and analytical procedures necessary for the design of a pellet drill bit are described. In the laboratory program approximately half the effort was devoted to the development of a better understanding of the physical process, and half to the empirical testing of new bit configurations. Factors such as pellet size and velocity, jet pump characteristics and modifications, and power distribution are discussed. The design and construction of a pellet impact bit for drilling a 9 in. hole with 1 1/4 in. steel pellets are described. Drilling performance in representative rocks, hydraulic operating characteristics, and metallurgical problems are discussed. INTRODUCTION This paper discusses an attempt to reduce drilling costs through the development of a drilling method basically different from rotary drilling. Publications on drilling research indicate that the major effort on the part of manufacturers and operators has been confined to research in rotary, or rotary-percussion drilling. This applied drilling research has produced significant gains in drilling rates and bit life, but thus far it has failed to achieve any major decrease in general drilling costs. From recent statistics released by Drilling Contractors, it can be shown that the total cost of drilling an "average" well in 1953 was about 2 1/4 times the corresponding cost in 1935. Analysis of the costs will show that they are not due to inefficient field operations; they are due instead to both an inflated economy and greater "average" well depths. Diligent work by the industry has reduced the backlog of potential economic improvements in rotary drilling to a group of rather marginal prospects. Increasing depths and poorer quality reservoirs at these depths continue to make the drilling cost problem more acute. Hence, the aim of the work reported here was to develop a new drilling method not limited by the inherent characteristics of rotary drilling and thus capable of attaining further significant reductions in drilling costs. The effort devoted to research on this proposed new drilling method was of considerable magnitude and represented .in excess of 37 man-years of research. The affiliates of Standard Oil Co. (N. I.) sponsored this development program which was conducted by the Drilling Methods Section of the Carter Research Laboratory in Tulsa, Okla. The purpose of this paper is to

Jan 1, 1957

By L. E. Young

The progress of mineral industry education will be limited to the period prior to World War II and will be considered as primarily a division of engineering education. Its relation to progress in the mineral industry and to the activities of the American Institute of Mining and Metallurgical Engineers will be emphasized. The history of mineral industry education has been recorded most completely and authoritatively by Dr. T. T. Read in his recent volume in the Seeley W. Mudd Series of American Institute of Mining and Metallurgical Engineers publications, entitled, The Development of Mineral Industry Education in the United States. Therefore, most of the historical details will be omitted from the discussion and an effort will be made to point out significant developments and to analyze trends. Introductory Statement and Significant Points in Progress Among the significant points in the development of mineral industry education in the United States since 1871, the following deserve particular mention: 1. The number of institutions offering undergraduate instruction has increased, particularly during the years 1890 to 1925; the number of students enrolled in undergraduate study has increased greatly, largely on account of the development of courses in petroleum engineering and increasing interest in metallurgy. 2. The selection of students to be admitted to undergraduate courses has been improved through the advisement, counseling, and guidance programs sponsored by the Engineers' Council for Professional Development. 3. Improved instruction methods all the training of the teaching staff have been promoted by the educational institutions themselves, but with the most effective stimulus and guidance of the American Society for Engineering Education and the Engineers' Council for Professional Development. 4. The accrediting of undergraduate curricula through the Engineers' Council for Professional Development, consitlering both qualitative and quantitative criteria, has done much to improve standards of instruction in engineering education for thc mineral industry. 5. Undergraduate curricula and courses are revised currently to meet the needs of the American mineral industry. New curricula are developed currently to permit specialization in technical fields, particularly in the fourth year of undergraduate work. 6. There has been a strengthening of the fundamental courses in mathematics, physics, mechanics, and chemistry, and at the same time these courses are being adapted to the most recent developments in the industry, including metallurgy, geophysics. petroleum engineering, petroleum technology, ceramics and SO forth. 7' There has been continuing improvement in plant and laboratory facilities of the educational institutions.

Jan 1, 1949

By D. F. Hewett

CONTENTS PAGE Introduction 1 Geologic features of eastern Oregon 2 Geology of the Sumpter quadrangle 2 Rocks older than the ore deposits 3 Rocks younger than the ore deposits 8 Lodes of the Sumpter quadrangle 8 Areal distribution 8 Geologic distribution 9 Major features of the veins 9 Mineralogy and texture of the veins 13 Ore deposits zonally disposed around the Bald Mountain batholith . . . . 13 Summary of mineralogy and texture 32 Zonal relations 33 Ore deposits zonally disposed around Greenhorn Mountain batholith . . . 36 Zonal relations 37 Relation of textural features to zonal arrangement of veins 37 Summary of zonal relations 38 Comparison with similar districts 39 Outlook for the district 40 INTRODUCTION IN an earlier report on the Sumpter quadrangle1 it was briefly stated that the quartz lodes of that region might be separated into several groups or belts on the basis of the kind and abundance of the minerals of which they are composed, and that these belts were disposed as areal zones around the principal intrusive mass of the region. When the opportunity to revisit the region was offered in 1929, as a result of a cooperative agreement between the State Mining Board of Oregon and the United States Geological Survey, it seemed advisable to pay particular attention to this feature of the lodes, in the hope that if the idea of zonal relations was confirmed it might serve as a guide in exploitation.

Jan 1, 1931

By David Cole

THE appearance of the Wilfley table in 1897 marked an epoch in the art of concentration of ores. The table has merited and received an almost unprecedented measure of public approval, lasting through its whole patent life. It has been very little improved in itself, or improved upon by competitors. The new machines bidding for popular favor have been of the Wilfley family, but without ability to do markedly better work than the parent m:-chine, or to (to more work except by the use of superimposed decks, the latter having obvious disadvantages. FIG. 1.-The- Butchart RIFFLE System. During the past three years there has been developed at Morenci a new type or arrangement of riffles applicable to the Wilfley type of concentrating table which corrects many of the objectionable features or limitations of the older system and obviates most of the difficulties encountered with it, particularly when handling ores having a low ratio of concentration. The new type is known as the Butchart1 riffle system, and its general arrangement is shown in Figs. 1 and 2.

Jan 2, 1915

By Ernest G. Enck

"The Rockets' Red Glare" of the Fourth of July is produced by strontium compounds. However, strontium has a number of less spectacular uses; i.e. in, or in connection with, the manufacture of ceramics, caustic soda, glass, railway and truck signal flares, beet sugar and steel. The principal ore of strontium for use both as the mineral and for production of compounds is celestite (SrSO4). This mineral varies in color from white to pink or blue depending on presence of impurities. It has a hardness of about 3.5 and specific gravity of 3.9. It is brittle and has perfect basal and prismatic cleavages. The high grade ore is obtained from England, hence it has become the standard for all other producing areas. (See analyses, Table 1.) The most objectionable impurity in English celestite is iron, coming mostly from adhering clay. On the other hand, Mexican celestite, while no higher in strontium sulfate content, is extremely low in iron (less than 0.2 pct), hence it is more desirable for certain uses. Most of the celestite in Spain and in the United States is high in calcium and barium, and commonly high in iron, and requires some form of beneficiation in order to be equal to English ore. The natural carbonate (SrCO3), or strontianite, is less important than celestite as a source of strontium, because it is less abundant and is not so widespread in its occurrence. Furthermore, it generally occurs with calcite (CaCO3) and lead sulfide (PbS), and for most uses these impurities are detrimental. This mineral is generally white or yellowish, forming needle-like crystals. It has a hardness of about 3.7 and a specific gravity of 3.68-3.8. The mineral is brittle and has an imperfect cleavage. A typical analysis of German strontianite is shown in Table 1. [ ] Distribution of Deposits UNITED STATES Occurrences of strontium minerals are known in Arkansas, Arizona, California, Illinois, Ohio, Pennsylvania, Tennessee, Texas, and Washington. Celestite occurs in the Vulture Mountains, Maricopa County, Arizona in beds three to ten feet thick associated with sandstone, gypsum, and barite. Because of the low strontium content, approximately 86 pct SrSO4, it has no commercial value. A large deposit of celestite occurs in San Bernardino County, California, in the Cady Mountains near Argus Station, about eight miles west of Ludlow. The celestite here is interbedded with shaly and clayey tuff in a zone about 300 ft thick. Massive, finely crystalline ore has been recovered with an analysis

Jan 1, 1960

By Edward H. Robie

EIGHT years ago in this department we had a column on the daiquiri cocktail, which, as we pointed out at that time, was invented by mining engineers. Since then we have said little or nothing about alcoholic drinks, but having set the precedent perhaps we may now pass along a little information about martinis. We have observed that as an aperitif the martini has become increasingly popular with men, so much so that we always feel a bit peculiar if we order something else. And the call is not for just a martini either; it must be a dry martini, and some of our friends are extremely particular on this point. The matter has achieved such importance that The New York Times devoted half a page to it a few Sundays back. Several recipes are given which may be useful to some of our members when entertaining visitors from this part of the country: The standard martini is said to be half gin, half Italian vermouth, two dashes orange bitters, and a pickled onion and twist of lemon peel. This progresses to the very dry martini, which calls for five parts gin and one part vermouth with an onion, olive, or lemon peel. Several alternatives are offered for the manufacture of the very, very dry martini. One is to pour a little vermouth into the mixing glass and pour it out again before mixing gin and ice. A second way is to put the ice in a sieve, hold it over the sink, and pour a little vermouth over it; then put the ice in the mixing glass. A third is to let the draft from an electric fan blow across an open vermouth bottle toward the mixing glass. Then the ultimate is the awfully dry martini, obtained by setting the vermouth bottle next to the mixing glass and turning the bottle slowly so that the label is exposed to the gin for perhaps a second. The Times notes that Mr. Martini himself personally prefers vermouth and soda.

Jan 1, 1952

By R. G. Gifkins

Mean boundary migration depended linearly upon creep extension under various conditions. Prolonged annealing or air-casting increased the proportion of immobile boundaries at a given stage of tests. Grain growth during annealing followed a power law. The linking of migration and sliding in creep and models for migration are examined; activation energies are calculated from these models. Segregation of oxygen to boundaries appears to be a controlling factor. THE idea that boundary migration and boundary sliding during creep are linked has often been suggested,'-5 especially when evidence has been obtained for an alternation between the two processes. In simple bicrystals sliding can occur without any migration6 but in more complex bicrystals7 it is observed. McLean' and Couling and Roberts 5 made measurements of both sliding and migration; McLean concluded that there was no linking of the two processes whereas Couling and Roberts supported the idea. McLean' also developed a model for migration during creep and annealing. The present work was initiated as part of a study of the recrystallization of lead during creep9 but has been extended to examine some aspects of the mechanism of migration. MATERIALS AND METHODS Specially refined high-purity lead* was used; the total of impurities, excluding oxygen, was 0.0008 wt pct. Oxygen was removed as far as possible by melting in vacuo, treating at 550" to 600°C with hydrogen and then casting at 450°C under a vacuum of 10-2 mm Hg; the oxygen was then believed to be 0.00005 wt pct. Specimens were made from extruded strip and then annealed at the required temperature in an air oven, controlled to 5 lac. They were then polished in a mixture of acetic acid and hydrogen peroxide. At this stage some specimens were etched to give grain contrast and their grain sizes determined by the intercept method. Migration of boundaries during creep was measured on the surface of polished (not etched) specimens, directly loaded at room temperature (22°C) or 50°C (52"C) to initial stresses of 300 to 700 1b per sq in. Twenty six vacuum-cast hydrogen-reduced specimens were used and, in addition, two specimens were from material air cast after hydrogen reduction ("partial" air cast) and two from air-cast material not treated with hydrogen ("full" air cast). Duplicates were made in the majority of tests. At intervals through the creep tests, specimens were unloaded and the migration measured, normal to the direction of migration at the point of measurement, for each boundary along standard traverses. Usually between 110 to 180 boundaries were measured at each of some twelve stages of the tests. Migration was measured on the projected image (X100) using prism illumination. Fig. 1 illustrates the method of measurement and shows the kind of variation of migration found from boundary to boundary and along particular boundaries. Fig. 2 gives an example of more extensive migration. The mean value of distance migrated was generally reproducible to *10 pct but occasional points gave a scatter of *20 pct. The correctibn applied by McLean8 to allow for the contribution sliding makes to the apparent distance of migration could not be used, for the sliding at most boundaries was too small to measure with the equipment available. There were two reasons for supposing the correction unnecessary in the present work;

Jan 1, 1960

By L. J. Kashar

By means of a continuum-mechanics approach, the following equation for calculuting the pressure for axisymmetric cold forging has been derived: where P is the extrusion pressure; K and n are the strength coefficient and the strain-hardening exponent of the steel, respectively; a and b are strain factors, primarily dependent on the die design; and R is the extrusion ratio. Data from compression and backward-extrusion tests performed at a crosshead speed of 0.5 in. per min on a nonwork-hardening material pure lead) and on a work-hardening material (1018 steel) indicate that the predicted extrusion pressures obtained from the derived equation agree very closely with the actual extrusion pressures. Therefore, the equation can be used to predict backward-extrusion pressures from standard mechanical-test data. The decision to cold-forge a part is based on several considerations, most of them related to economics. However, the production feasibility is usually determined by the pressure needed to extrude the part and the availability of tooling that can repeatedly sustain this pressure for over several hundred thousand cycles. A major contribution to the cold forging of steel would result from the development of steels that could be forged at reduced pressures and from the development of techniques that permit the accurate prediction of the pressure necessary to cold-forge the part. The ability to predict extrusion pressures is necessary to properly design the tooling required for the cold-forging process. Attempts to theoretically predict the cold-forging pressure, especially for work-hardening materials, have been unsuccessful in previous investigations of the cold-forging process.1-3 Many empirical relations have, therefore, been proposed relating the extrusion pressure to the chemical composition,' the extrusion ratio,2 the tensile strength,273 and hard- ness, 2 and the yield strength3 of the material being extruded. However, these empirical relations are valid only for the specific conditions used in their development, and, in some cases, the empirical relations had a scatter band greater than * 10 pct. A study of the extrusion process was undertaken with the objective of developing predictor equations for extrusion pressure by using a continuum-mechanics approach to the problem of metal flow. The derivation of such an equation is summarized in this paper, and data are presented that confirm the validity of the derived equation. DERIVATION OF THE EXTRUSION EQUATION If the extrusion of a material through a die is considered in terms of a work balance, it has been shown that it is possible to derive an equation relating the work necessary to move the material through the die to the deformation work undergone by the material. The work done, W, in moving the material is given by where F is the extrusion force, I, is the distance the material was moved, A, is the original cross-sectional area of the material, P is the extrusion pressure acting on A,, Ap is the cross-sectional area of the extrusion punch, and Pp is the punch pressure acting on Ap. The deformation work, WD, is given by where V is the volume of metal deformed, ay is the flow stress of the metal, and E is the total strain undergone in the extrusion. If there are no frictional losses, the work done, W, is equal to the deformation work, WD; therefore, Since the volume of a metal does not change during deformation,

Jan 1, 1968

By R. W. Woodward, G. K. Burgess

minum bronze. Nearly all copper-base alloys are improved by rolling processes, but the copper-aluminum alloys seem to possess equally good properties when cast or rolled; this is a remarkable metallurgical fact. Another important property of these copper-aluminum alloys is their resistance to alternating stress. Many tests indicate that their resistance is greater in this respect than that of some steels, and I have seen instances when cast aluminum-bronze bolts have outlived five steel bolts in foundation work subject to severe shocks. I mention these various instances to indicate that work originally started as research for substitution of one material for another frequently develops 212 article which has properties not possessed by the original metal or alloy. I have dwelt particularly on the aluminum bronzes because recently I have done more special work on them than on other alloys, but I believe that the use of aluminum itself, in many combinations of metals, is a very important subject for investigation. Undoubtedly, after the war, the cost of aluminum will be reduced from its present price, and considering its low specific gravity it offers a very interesting and important field for research in the non-ferrous business. Naturally, if combinations containing aluminum can be developed in view of a probably increasing supply of metal, the cost will be reduced. This will benefit the industry generally and will immediately conserve tin. Bronzes, Bearing Metals, and Solders G. K. Burgess* and R. W. Woodward,† Washington, D. C.— From a metallurgical standpoint, there are several ways in which a reduction of the tin consumed in commercial non-ferrous and white-metal alloys can be effected. First, a reduction of the tin content of the alloy; second, substitution of part or all of the tin content by some other metal; third, a substitution of a different type of alloy, which in some cases also involves a change in mechanical design. The Bureau of Standards has been studying these methods of conservation for tin alloys, particularly in regard to babbitts and bearing metals, bronzes, and solders. Much of the information sewed by the Bureau was obtained from answers to questionnaires sent to manufacturers and users of these materials, so that, in general, any of the following suggestions or recommendations can be considered as being practical and as having already been thoroughly tried. Bearing Metals There is no question that the tin content of nearly all bearing metals can be reduced to some extent, and in some cases actually eliminated

Jan 1, 1919

By E. H. Rose

THE present symposium on flotation machines provides a unique opportunity for group reappraisal of standard or habitual viewpoints. A symposium is by definition a trading post for observations that might prove complementary when put together, thus leading to useful thoughts that might not have been evident from the original observations individually. The present paper, pertaining to power consumption in flotation, is deliberately an experiment of that kind. Data were solicited from the operating superintendents of several of the major flotation plants on this continent. The information received, covering a total of 14 separate operations in seven plants, has been collated in terms of power intensity; that is, the total power input per cubic foot of cell volume. This departs somewhat from the conventional expression of power per ton of ore treated, though the latter figure is also given for the plants concerned. Because of wartime restrictions, particularly in Canada, identifying names of plants and machines are withheld. Not as a firm conclusion, but as a tender for symposium discussion, the following points are made: I. That flotation is a power operation in much the same sense as ball milling. 2. That contact time is the complement of power intensity and may be decreased as the latter is increased. 3. That a decrease in flotation power per ton of ore can result from an increase in power intensity. ANALOGY WITH GRINDING In grinding, it has become axiomatic that the more the power put into a given operation, the greater the work performance. The analogous view is now proposed that the effectiveness of any flotation operation will be related to the mechanical energy usefully applied to that operation. Under this view, each increment of properly directed input energy should give either an increase in unit tonnage or an improvement in the metallurgical separation (concentrate grade or tailing grade), until either the theoretical limit imposed by mineral constitution or surface condition is reached, or until further increments of power become uneconomic under the law of diminishing returns. Thus, as in grinding, there are two forms of output that greater power input or improved utilization of power can produce. As for the law of diminishing returns, lower unit cost of power through improved generating efficiency in the past decade has pushed the limit back. Over and above this, however, there is strong indication in the operationg data herein presented that higher power intensity results in an actual economy of over-all power consumption. Like crushing and grinding, flotation accomplishes a rupture of the forces of molecular attraction, by taking particles from the common environment and placing them in the two separate environments of

Jan 1, 1944

The earliest U. S. patent on a crushing machine was issued in 1830. The device incorporated the drop hammer principle later used in the famous stamp mill, whose history is so intimately linked with that of the golden age of American Mining. Eli Blake invented the first successful mechanical rock-breaker --the Blake jaw crusher--patented in 1858. The idea of his mechanical principle, the powerful-toggle linkage, was so sound that today the Blake -type jaw crusher is the standard by which all jaw crushers are judged. In 1881, Philetus Gates was granted a patent on a machine which included in its design all the essential features of the modern gyra tory crusher. For some years after these pioneer machines were developed, crushing requirements were quite simple. All underground mining and open-pit quarrying was done by hand with tonnages and rock sizes generally small. In the milling of precious metal ores, stamp mills were popular as the final reduction machine. The stamps were fed with an ore size readily produced by one break through small gyratory or jaw crushers. Even in large underground mining operations, there was no demand for large crushers; and increasing tonnage requirements were met by duplicating the small units. The Homestake gold mine in South Dakota was a good example, where in 19 15 twenty -two small Gates gyratory crushers were used to prepare the ore for the batteries of 2,500 stamps. Open pit mines and the introduction of the steam shovel created a demand for larger crushers, which led to the development of a gyratory with an 18 in. (457 mm) receiving opening. Up to this time, the jaw crusher had kept pace with the gyratory, both from the standpoint of receiving opening and capacity; but now the gyratory stepped into a 15-year lead with larger and larger sizes. This trend toward large primary crushers started just before the turn of the century, and in 1910 crushers with 48 in. (1.22 m) receiving openings were being built. About this time, the jaw crusher was revived in a great contribution to primary crushing by the introduction of the 84 in. by 60 in. (2. 13 by 1. 52 m) machine. This crusher was usually followed by a 24 in. (6 10 mm) opening gyratory crusher for secondary breakage, a system which reawakened the industry to the use of the jaw crusher as a primary breaker. It is interesting that although his machines never came into general use, Thomas A. Edison ranks as a pioneer in the development of the large primary breaker. Edis on reasoned that the recoverable electrical energy

Jan 1, 1978

By E. H. Rose

The present symposium on flotation machines provides a unique opportunity for group reappraisal of standard or habitual viewpoints. A symposium is by definition a trading post- for observations that might prove complementary when put together, thus leading to useful thoughts that might not have been evident from the original observations individually. The present paper, pertaining to power consumption in flotation, is deliberately an experiment of that kind. Data were solicited from the operating superintendents of several of the major flotation plants on this continent. The information received, covering a total of 14 separate operations in seven plants, has been collated in terms of power intensity; that is, the total power input per cubic foot of cell volume. This departs somewhat from the conventional expression of power per ton of ore treated, though the latter figure is also given for the plants concerned. Because of wartime restrictions, particularly in Canada, identifying names of plants and machines are withheld. Not as a firm conclusion, but as a tender for symposium discussion, the following points are made: 1. That flotation is a power operation in much the same sense as ball milling. 2. That contact time is the complement of power intensity and may be decreased as the latter is increased. 3. That a decrease in flotation power per ton of ore can result from an increase in power intensity. Analogy with Grinding In grinding, it has become axiomatic that the more the power put into a given operation, the greater the work performance. The analogous view is now proposed that the effectiveness of any flotation operation will be related to the mechanical energy usefully applied to that operation. Under this view, each increment of properly directed input energy should give either an increase in unit tonnage or an improvement in the metallurgical separation (concentrate grade or tailing grade), until either the theoretical limit imposed by mineral constitution or surface condition is reached, or until further increments of power become uneconomic under the law of diminishing returns. Thus, as in grinding, there are two forms of output that greater power input or improved utilization of power can produce. As for the law of diminishing returns, lower unit cost of power through improved generating efficiency in the past decade has pushed the limit back. Over and above this, however, there is strong indication in the operationg data herein presented that higher power intensity results in an actual economy of over-all power consumption. Like crushing and grinding, flotation accomplishes a rupture of the forces of molecular attraction, by taking particles from the common environment and placing them in the two separate environments of

Jan 1, 1947

By E. H. Rose

The present symposium on flotation machines provides a unique opportunity for group reappraisal of standard or habitual viewpoints. A symposium is by definition a trading post- for observations that might prove complementary when put together, thus leading to useful thoughts that might not have been evident from the original observations individually. The present paper, pertaining to power consumption in flotation, is deliberately an experiment of that kind. Data were solicited from the operating superintendents of several of the major flotation plants on this continent. The information received, covering a total of 14 separate operations in seven plants, has been collated in terms of power intensity; that is, the total power input per cubic foot of cell volume. This departs somewhat from the conventional expression of power per ton of ore treated, though the latter figure is also given for the plants concerned. Because of wartime restrictions, particularly in Canada, identifying names of plants and machines are withheld. Not as a firm conclusion, but as a tender for symposium discussion, the following points are made: 1. That flotation is a power operation in much the same sense as ball milling. 2. That contact time is the complement of power intensity and may be decreased as the latter is increased. 3. That a decrease in flotation power per ton of ore can result from an increase in power intensity. Analogy with Grinding In grinding, it has become axiomatic that the more the power put into a given operation, the greater the work performance. The analogous view is now proposed that the effectiveness of any flotation operation will be related to the mechanical energy usefully applied to that operation. Under this view, each increment of properly directed input energy should give either an increase in unit tonnage or an improvement in the metallurgical separation (concentrate grade or tailing grade), until either the theoretical limit imposed by mineral constitution or surface condition is reached, or until further increments of power become uneconomic under the law of diminishing returns. Thus, as in grinding, there are two forms of output that greater power input or improved utilization of power can produce. As for the law of diminishing returns, lower unit cost of power through improved generating efficiency in the past decade has pushed the limit back. Over and above this, however, there is strong indication in the operationg data herein presented that higher power intensity results in an actual economy of over-all power consumption. Like crushing and grinding, flotation accomplishes a rupture of the forces of molecular attraction, by taking particles from the common environment and placing them in the two separate environments of

Jan 1, 1947

By J. D. Bateman

At Giant Yellowknife, where high grade gold-bearing orebodies are highly irregular in shape, geology has been applied extensively to the mining of ore. The classical functions of the mine geologist in the fields of exploration and mine development have been extended to guide ore extraction, ensuring "clean" mining and effectively reducing waste dilution. THE property of Giant Yellowknife Gold M.ines Ltd. is situated west of Yellowknife Bay on the north shore of Great Slave Lake, a distance of 600 air miles north of Edmonton, Alberta. The Giant claims were staked in 1935, the company was formed in 1937, and the main orebody system was disclosed by diamond drilling in 1944 following a geological study of the property by A. S. Dadson,* Consulting Geologist for the company. Production began in 1948 at the rate of 200 tons per day and, during 1950, reached a daily rate of 425 tons. During the first 3 years of operation a total of 366,000 tons was milled with an average grade of 0.79 oz of gold per ton. Work is in progress with an expansion to 700 tons per day in view. A descriptive account of the geology and gold-bearing shear zone has appeared previously.* The rock formations in the vicinity of Yellowknife Bay have been subjected to protracted pre-Cambrian tectonic deformation culminating in a series of late faults having a cumulative horizontal displacement exceeding 11 miles. The Giant property is underlain by part of an Archean sequence, several miles thick, consisting of basic volcanic flows and minor intercalated tuffs. The volcanic succession forms the west limb of a major syncline, the flows facing east, but overturned on Giant property to dip west at 65" to 75". Orebodies are confined to shear zones up to 200 ft in width, which were formed along early thrust faults. The shear zones assume fold-like attitudes, the larger of which have an amplitude of several hundred feet. The rock formations beyond the limits of the zone of shearing do not reflect the simulated folds, the axes of which are within a few degrees of the strike of the flows. The schistosity and most of the planar elements in both the shear zones and the orebodies dip west • A. S. Dadson and J. D. Bateman: Structural Geology of Camdian Ore Deposits, Can. Inst. Min. Met. Jubilee Volume (1948), PP 273-283. at angles between 65" and 75", generally corresponding to the dip of the flows. The planar elements within the shear zone system thus dip more or less constantly west whether the shear zone is flat, vertical, or expressed as east or west dipping limbs. The shear zones reflect the deformation and alteration of the basic volcanic flows into chlorite schists which, in most places, have undergone metasomatic replacement to form chlorite-sericite-carbonate schists or sericite schists. The boundaries between the shear zones and country rock, although often gradational, usually can be defined within a few feet or even inches as they are expressed by the limits of metasomatic alteration. Orebodies may occupy a small or large proportion of the shear zone and, although they generally conform to the shape of the zone, their morphology is much more complex. Ore boundaries in some instances are sharp and can be delineated with a chalk line; but more generally, a large proportion of the ore boundaries is not visually obvious and can be determined only by the perception acquired by the geological mapping of ore or study of drill cores. Ore shoots in fold-like attitudes may transect the planar elements of the shear zone at any angle; yet the schistosity within the ore shoot may be coincident with that in the enclosing shear zone. Thus it is clear that problems may arise in the delineation of mining boundaries. Ore generally consists of 20 pct or more quartz with ferruginous carbonates in sericite schist deposited in two dominant stages. The earlier stage limits of quartz with carbonate, pyrite, and very fine-grained arsenopyrite in lenses and bands. The later stage consists of quartz-carbonate lodes, in

Jan 1, 1952

By J. C. Allen, E. V. Herring

CHILE EXPLORATION CO's. oxide plant at Chuquicamata, the largest of its kind, started in 1915 with initial operations at the rate of 10,000 tons of ore per day. Anaconda Copper Mining Co. acquired the controlling interest in 1923 and in the next few years modernized the plant and mine, extending operations to a nominal capacity of 490 million lb of electrolytic copper yearly by 1942. Plant production averaging 495.7 million lb of copper per year was maintained for the eight-year period extending from 1941 to 1948 inclusive. During the World War II production reached its highest level when in 1944 the output was 531.8 million lb. The maximum production for any one month was 49.6 million lb, achieved in May 1947. E. A. Cappelen Smith, consulting metallurgist for M. Guggenheim's Sons, worked out the first process for the treatment of Chuquicamata copper oxide ore about 1913, and directed a staff of 20 young engineers operating a pilot plant at Perth Amboy, N. J., on three shifts for an entire year. The copper is extracted from the oxide ore by a hydro-metallurgical treatment described in detail by T. A. Campbell in AIME Transactions, Vol. 106, Copper Metallurgy, 1933. The Oxide Plant is divided into five main operations as follows: (1) Crushing, (2) Leaching, including dechloridizing and sulphur dioxide treatment, (3) Tailing disposal, (4) Electrolytic precipitation, (5) Smelting and melting. Crushing The present Primary Crushing Plant has been in operation 25 years, during which time 4,342,112 cars of 70 tons net capacity have passed through the dumpers. The ore as received from the mine is frequently blocky and up to 5 feet in section. It is reduced to 9-in. size by two 60-in. gyratory crushers, and transported to the east and west ore bins by 60-in. conveyors. In the secondary crushing operation the ore is first reduced to 2 to 3-in. size by five standard 7-ft cone crushers and two No. 10 gyratories. This product is then reduced to 3/8-in. size by fifty-two 48-in. vertical Symons disc crushers and is transported by belt conveyors to the loading bridges from which it is charged° into the leaching vats. Conveyor belt No. 14 which serves the eight vats in the east section of the leaching plant has transported over 125 million tons of ore, which is a record for any type of conveyor. This 60-in. belt travels at a speed of 600 fpm and carries loads up to 5500 tons per hr. Leaching The ore is leached in 14 vats, averaging 13,500 short tons net capacity each. These vats are lined with a natural mastic which adequately resists the sulphuric acid leaching solutions. The leaching process as presently conducted comprises soaking the ore with successive volumes of leaching solution and spent electrolyte and withdrawing successive volumes of enriched solution from the vat to a total of approximately 3.2 million gal. The first 1.6 million gal of this volume withdrawn is called strong solution and contains about 35 parts of copper per 1000. The balance of 1.6 million gal subsequently withdrawn is stored in a

Jan 1, 1952

By Louis W. Huber

The Liberty tunnels extend through a very steep hill in Pittsburgh (locally called Mount Washington) for a distance of slightly over a mile. The two tunnels parallel each other and are 59 ft. apart, center to center (Fig. 1) They are for vehicular traffic, and in each there is ample space for two lines of vehicles. Northbound traffic, toward downtown Pitts- burgh, is confined to the east tunnel and southbound traffic to the west tunnel. The South Hills residential district is separated from downtown Pittsburgh by Mount Washington, a steeply rising escarpment south of and adjacent to the Monongahela River. The Mount Washington district immediately south of the Monongahela River has been reached, for many

Jan 1, 1927