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By Gerhard Derge

The rapidity of oxygen steelmaking has created renewed interesl in the mechanisms and kinetics of the various processes involved. These studies can now be approached with increased confidence because of the more adequate thermodynamics, physical chernislry, and relevant engineering infornzation available. These combine to offer promise of useful control and automation if the details of each specific problem can be sufficiently well-defined. The hot filament microscope provides new insig-hi into many of these problems by direct observation at temperature, allowing the appraisal of the relative importance of thermal gradients, convection, and intermediate phases. All of these have a direct bearing on the overall rates and control of most pyrometallurgical processes. It is one of the special privileges of the Howe Lecturer to become better acquainted with this distinguished man. All of us interested in the metallurgy of iron and steel have developed some sense of his widespread influence by the annual reference to his work in these Memorial Lectures. Each lecturer since 1923 has been able to identify his personal specialty with the broad interests of Professor Howe. This has become even more rewarding with the background information from Professor Gensamer's monograph, "Henry Marion Howe's Contributions to Metallurgy, which brings out his happy combination of intellectual and professional activities. Beyond this, I found my personal bond with Professor Howe in his book on The Metallurgy of Sleel first published in 1891.' After a detailed discussion of the dephosphorizing action of slags, he relates the mineral species identified in the frozen sample to the chemistry of the fused slag with the question, "Is it not most philosophic to hold that these independent entities were wholly integrated in the initial magma; that their presence in it was wholly

Jan 1, 1968

By Ben L. Seegmiller

Introduction The profitability of an open pit operation depends to a large extent on the use of the steepest pit slopes possible, provided they do not fail during the life of the mine Optimum pit slopes can be designed by using rock mechanics technology. Such technology includes gathering and analyzing pertinent data prior to pit design, rigorous application of rock mechanics principles during the design process, remedial actions to improve slope stability, and monitoring slopes during the life of the mine. Ignoring rock mechanics technology in pit design may place the entire mining enter- prise in jeopardy Disruption of mining operations and safety hazards caused by slope instability can be severe. A classic case in point is the major slope failure (Seegmiller, 1972) that occurred at a southwestern US copper mining operation in 1970-1971 Minor slope stability problems initially amounted to no more than a nuisance, but in time, the unstable zones became larger, causing greater disruption of production and threatening safety Mining at or near the slide toe had continued throughout the period of slope failure, aggravating the already unstable condition The original plan for this pit called for mining the overall pit slope to 37.5" Mine operators were forced to lessen the back- slope angle in the hope that the failing rock mass would stabilize. However, stabilization did not take place, and when mining in the area of failure was completed, the overall slope had been reduced to approximately 27" Mining had been successful in removing only a small portion of the ore body at a high stripping ratio.

Jan 1, 1979

By Joseph P. Lyden

In the Picher field, structure made openings for the circulation of the mineralizing solutions by flexing, shearing, and fracturing the sedimentary beds. This structure is used with the spatial and genetic relationship of the ore minerals, sphalerite and galena, and the gangue minerals, dolomite and jasperoid, in prospecting for ore bodies.

Jan 12, 1950

By Willard Ide Pierce

THE native methods of mining and milling gold-ores in Korea may not present any specially new features, but are chiefly interesting as primitive practices, still followed at the present day. In mining where the rock is hard, wood is piled up near the face of the tunnel or drift and then set on fire. The fire bums for twenty-four hours, after which, three or four days are given for the rocks to cool, before the miners go in and extract the loosened material. The only tools in use are a hammer, a steel gad about five inches long, and a flat spoon-shaped basket. The pay-streak or chute is closely followed, and in places the workings are very small and intricate. Rock not yielding more than $6 per ton is not extracted. The manner of opening a mine is to begin on the outcrop, where the rock pays, sink a shaft or run a tunnel on the vein, and follow the pay-streak and more easily worked ore. When this is exhausted, or the influx of water interferes, the opening is abandoned and operations are begun again at a new spot on the outcrop. The place usually selected is as high up on the mountain as possible, and in no instance is advantage taken of the topography by running a cross-tunnel to tap the vein. The ore is brought in baskets to the surface, and there placed in sacks, holding about 150 1bs. each, and taken to be crushed. The rock is first crushed between two stones, as shown in Fig. 1. The upper stone is oval and rounded on the bottom, so that it can be readily rocked on the under stone. Two pieces of wood, used for handles, are placed, one on each side, in grooves in the upper stone, and held in position by means of straw rope. The lower stone is not smooth on top, but exhibits three elevations running lengthwise, with shallow depressions between them. The middle elevation acts as a fulcrum on which the upper stone may be rocked. From two to ten men, according to the weight and size of the stone used, are employed in rocking. With four men, from 300 to 350 lbs. of rock can be crushed per day, the stone making 50 complete oscillations per minute. The men squat on each side of the stone and move it to and fro by means of the wooden handles. The ore is fed, in pieces up to the size of an egg or even larger, by one of the men. During the crushing one man on each side, as the stone rocks from

Jan 1, 1890

By G. S. Cole, G. F. Bolling

Recent experitnents on crystal growth have indicated that thermal convection in the liquid influences the conditions of solidification and subsequent crystal characteristics. However, the important columnar-equiaxed transition in castings has not been examined as a function of this free convection. An experiment is presented here to demonstrate a change in the transition as convection is suppressed. Lead- and tin-based alloys were cast in stainless-steel molds either without special procedures or with thin grids placed horizontally across the tnold diameter in portions of the castings. While the grids allowed some diffusion of heat and mass transport to occur during solidification, massive fluid flow was hindered in the vegiotz of the grids. A larger percentage of columnar grains appeared in the grid region than in other parts of the special castings or than in all regions of normal castings. This structural difference can be explained in terms of thermal convection. THERE is little information about the possible effects of natural convection which may occur during the solidification of a casting. Ruddlel-3 has stated that the movement of liquid metal produced by convection might assist in the transfer of heat, but in the absence of confirmatory evidence he suggested that this idea be treated with caution. chalmers4 considers, but does not assess, the importance of fluid motion in casting, while Tiller5 suggests that, in liquid metals which have extremely high thermal diffusivities, the effects of thermal convection may be negligible. In addition, throughout the literature, there are statements indicating the possible presence of thermal convection but without proofs or evaluations of its influence. Natural convection may be divided into two categories: thermal convection, where liquid flow is due to density differences caused by temperature gradients in the liquid (e.g., liquid near a cold wall falling with respect to hotter liquid further away from the wall); and solute convection, where flow is due to density differences as a result of concentration gradients, see Wagner.6 In most instances thermal convection predominates over solute convection. Nevertheless, even thermal convection has not been studied in any detail and is difficult to handle, since it can vary quantitatively from one experiment to another depending upon the exact positions of heat sources and sinks or the attitude of the container with respect to gravity. In spite of these difficulties it is possible to realize that a higher rate of heat transfer must result with convection than in any situation where conduction acts alone. Thus, the degree of increased heat transfer is the important undecided question in considerations of casting. It will be demonstrated here that a reduction in convective flow leads to a suppression of the columnar-equiaxed transition in a casting and that this suppression is related to a decrease in the degree of heat transfer. Fortunately it is possible to do this by comparative rather than quantitative measures. The insertion of horizontal plates or grids into a simple mold does not appreciably influence heat transfer per se, but in reducing the allowable height through which colder liquid may fall freely it reduces natural thermal convection. This leads to a reduced fluid velocity, a different flow pattern, and an altered heat transfer which can be measured. EXPERIMENTS The alloys were prepared from elements of 99.9 pct nominal purity and cast into an uncoated 2-in.-diameter stainless-steel mold insulated at the bottom with a suitable refractory so as to ensure predominantly radial heat flow during freezing. Casting was performed either in empty molds or in molds containing grids or plates in various configurations . The grids were constructed of aluminum screen (1/32 in. thick) which was cut to the desired shape, cleaned in aqua regia, and coated with a mixture of 2 parts aqua dag graphite -1 part Sauereisen cement mixture, thus supplying a hard, protective, and insulating layer approximately 0.006 in. thick. The

Jan 1, 1965

By William Grady

FACTORS entering into the market value of coal are its grade, and the cost of labor, material, and capital. Reduction in these costs cannot be expected in the future, and it therefore follows that greater economy in their use must be accomplished if it is desired to hold the present markets; more tons of product per unit must be obtained in a given time from both men and material, and means devised for increasing the percentage of lump and domestic sizes. This paper is presented for the purpose of inviting discussion and stimulating thought in regard to the influence of methods of procedure and plans of mining, upon the quality of coal, the economical use of labor, material, and capital, and the cost of production. Even the better methods of the. present day leave much to be desired, and a review of the reasons and necessities for their use may be of value in pointing the way to more economical mining. The writer does not attempt to discuss untried methods of mining, but rather to state clearly and concisely a summary of the results obtained under the several methods of procedure that have been adopted. In some instances a high degree of concentration has been effected, and comparisons of the results obtained with the results of lesser degrees of concentration form an interesting study. It is believed that much may be accomplished in improving the quality, and in the more economical use of labor, material, and capital, by concentration, and it is to concentration in particular that your study is invited. The writer, who .has occasion to visit many mines, most of which ire under different management, has been inclined to the opinion that improvement has been retarded in many ways and that mining has not as yet been entirely freed from the early-day practices which were forced upon it.

Jan 5, 1915

By M. N. Dunlap

This article summarizes the successful incorporation of a flash-flooding stream into the tailing-disposal system at the St. Joseph Lead Company's Federal Division mill, in St. Francois County, Missouri. The mill was built in 1906, by the Federal Lead Co., as its No. 3 mill for the concentration of southeastern Missouri lead ores, and damming Davis Creek was first considered in 1908. Plans were made also in 1925 and in 1937. Dam layouts varied from a reinforced concrete structure to a barrier built of tailings alone. .All were abandoned because of liability that would be incurred by flash-flood rupturing of an incompleted dam. By 1942 mill capacity had reached 12,000 tons daily and increasing tailing-disposal costs again focused attention on the Davis Creek basin. This time the problem of riskless dam building was solved. Moreover, most of the items that appeared in the answer were materials that normally required wasting and the major portion of labor involved performed functions that it mould normally have done, without constructive benefit, elsewhere. The new layout also made possible advantageous changes in mill-water supply pumping; static head being reduced from 514 ft to 9; ft, friction head from 4900 ft to 2860 ft, and pumping can be done now at periods of low power demand. The following facts about the Davis Creek watershed explain the 36-year reluctance to reach these, and other, benefits. Size: 3600 acres. Outline: roughly pear-shaped, 21/2 by 3 miles at the largest dimensions. Topography: mostly hill land with steep valleys. Relief: 760 minimum to 1153 maximum. Maximum rainfall: 3.36 in. per hour. Permanent water: none natural. Now 2200 gpm discharged through 16-in. drill hole from mine pumps underground. Geology: the 17 acres near dam centerline is Bonne Terre dolomite covered with sand and gravel eroded from Pleistocene beds that cap some of the hills. The remainder of the pond basin is on Davis shale. Soil cover: mostly second-growth timber with some abandoned fields; all subject to careless burning, which reduces water-retaining ability. Flood history: flash-flooding that had taken at least one life. Downstream values: two railroad bridges before creek intersection with Flat River. Immediately below confluence, a highway bridge, a low-water crossing, a few business establishments and several blocks of residences. In July 1942 the following segments of an ingenious plan were put into operation. A ditch 270 ft long, 16 ft wide and 51/2 ft deep was excavated to bedrock in the gravelly soil along the proposed dam's centerline. This excavation, which exposed many gravel-filled channels, was run full of slimes drawn from a riser set in the top of an adjacent tailings-disposal

Jan 1, 1949

By E. H. Lovitt, A. C. Skerl

After nine years of continuous operation it is a good idea to take stock of the geological concepts that guide the working of a mine. This is a welcome opportunity to discuss the Lovitt gold mine, which has been a challenge in both the geological and operational fields. The occurrence belongs to the extensive group of epithermal gold-silver deposits once worked so vigorously in the western states, Hungary, and New Zealand and still producing in Central America and the Philippines. These deposits are characterized by a relatively shallow vertical range up to 2000 ft, and often much less, and by bonanza orebodies of even more limited range. The area in which they occur is frequently widespread, so that exploration must be directed along and across the strike of known favorable structures in search of repetitions.

Jan 9, 1958

By S. F. Reiter

The paper presents isothermal recrystallization curves for 0.08 and 0.15 pct C steel at subcritical temperatures following small amounts of plastic deformation. The effects of deformation, temperature, and aging on nucleation and growth rates ore described. The free energy of activation for grain boundary migration in steel is given. SEVERAL excellent reviews of the literature have appeared concerning the recrystallization of metals.'-' The present investigation follows the approach advanced by Mehl, Stanley, and Anderson,6-7 in which the rate of recrystallization was analyzed in terms of N, the rate of nucleation, and G, the rate of growth of recrystallization nuclei. Two lots of low carbon, capped steel of the analysis given in Table I were studied. Each lot consisted of a 150 lb coil which had been hot rolled to 0.083 in. and then cold rolled to 0.042 in. at the mill. Strips 0.930 in. wide were sheared perpendicular to the rolling direction. Both steels were normalized before studying their recrystallization characteristics. The strips were cleaned, painted with a magnesia-acetone paste, and made into packs of equal weight, wrapped in 0.002 in. copper foil. The packs were placed in a salt bath at 900°C for 30 min and air cooled. A relief anneal followed in a second salt bath for 15 min at 650°C. The relief anneal was found necessary from early tests in which a longer incubation period and slower rate of recrystallization were observed in relief-annealed lot A steel than in similar material which was strained and recrystallized directly after being normalized. This effect, which indicates the presence of transformation and/or cooling stresses in steel air cooled from above the A, temperature, has also been observed by Samuels8 and Masing.9 Figs. 1 and 2 show the microstructure of lot A and B materials and illustrate the rather uniform No. 8 ASTM grain size produced by this heat treatment. Winlock and Leiter10 observed that strip specimens which had their sharp edges removed elongated more uniformly than those which were not polished. Similarly, when the sheared edges were removed on a belt grinder, it was found in the present investigation that such samples recrystallized more uniformly than did unpolished strips. Therefore, all strips were carefully rounded prior to their extension. The approximate strain limits for the production of large recrystallized grains are from 6 to 12 pct extension." It was found that for the purpose of this investigation, 8 and 9 pct elongation were suitable deformations. The strain rate employed was 0.01 in. per in. per min and produced a yield point elongation of 4 pct. Winlock and Leiter found that mild steel of No. 8 ASTM grain size gave the same yield point elongation when extended at 0.012 in. per in. per min. All of the lot A and B strips extended in tension developed a straight, stretcher strain line at each grip when the upper yield point was reached. The lines were parallel and made an angle of 55" with the edge of the strip. They approached each other with increasing strain and met near the center of the sample at the end of the yield point elongation. Immediately thereafter, a small drop in load was observed and then the load increased in a regular manner with increasing extension. The grips were initially 8 in. apart. After extension, the 6 in. gage length was carefully cut into 1 in. samples. The remainder of the strip was discarded. After a flash pickle in hot 50-50 hydrochloric acid, six samples, each of which had been taken from a different strip, were placed in a basket and submerged in a lead pot for isothermal recrystallization. Although no recovery effect was observed, strain aging did occur after extension. Therefore, samples were always recrystallized within 24 hr after their cold deformation. After recrystallization, the samples were etched with a solution comprised of one part by volume of nitric acid with three parts of water. Bromide printing paper was exposed directly at low magnifications and later used with a mask to measure the desired quantities. First, the average diameter of the largest grain visible in each sample was determined using dividers. Next, the number of recrystallized grains per unit area was counted and recorded as n. Then, for each sample, the combined area of the recrystallized grains was measured by transcribing the grain outlines to standard graph paper. Many determinations of the area of the recrystallized grains were repeated five times and indicated a standard error that was not greater than 25 pct. The average area for six samples was divided by the area of the mask to yield the percentage recrystallized. Recrystallization of 0.08 Pct C Steel The progress of recrystallization at 670°C after 8 pct elongation of lot A steel is shown in Fig. 3, a through f. The shapes of the growing crystals are approximately equiaxed, as is assumed in the

Jan 1, 1953

By L. F. Wright

Operations of these two mining companies have some superficial similarities-geographical location, production rates, comparatively large ore reserves; but any critical comparison of mining and milling procedures or of economics must take into account the striking differences in geological environment and mineralogy of the respective ore deposits. The properties are only about 20 miles apart on an upland surface of moderate relief (part of the Interior Plateau of central B. C.) which occupies an area east of Thompson River between the towns of Ashcroft and Merritt. Being in the "dry belt" of British Columbia, water conservation is necessary at both operations.

Jan 12, 1963

By AIME AIME

THE Wednesday morning session was devoted entirely to the consideration of the tentative code for coal mine ventilation. A. W. Hesse is chairman of this subcommittee. E. A. Holbrook presided at the session and assisted by A. W. Hesse and George S. Rice, gave the historical background of the movement for a ventilation code. They stressed the point that this was to be an engineer's code, and that it was hoped that it might serve as a valuable guide to operating companies in securing greater efficiency as well as safety in ventilation. The discussion was continued into the afternoon session, and by adjournment the entire' code had been covered.' Many sections were referred back to the Committee for further study and several new sections were suggested. The various opinions and new ideas brought out during the discussion proved very valuable and. contributed much to the progress being made in the preparation of this code.

Jan 1, 1929

GENERAL Broken Hill lies at latitude 31º58'S longitude 141º27'E on the semi-arid western plains of New South Wales. From discovery in 1883 until the end of 1969 the lead-silver-zinc orebody has yielded 107,570,000 tons of ore for an overall grade of 13.1% Pb, 6.9 oz Ag/ton and 9.4% Zn.* The orebody currently supports the activities of four mines which are, from north to south, North Broken Hill Limited, Broken Hill South Limited, The Zinc Corporation, Limited and New Broken Hill Consolidated Limited. It is estimated that in 1969 these mines produced about 8 ½ % of world lead production and about 6% of world zinc production. In 1905 The Zinc Corporation, Limited was founded to recover zinc from tailings dumps and moved into mining in 1911 on leases purchased from Broken Hill South Limited at the then southern extremity of the "line of lode". New Broken Hill Consolidated Limited was formed in 1936 to explore and mine for deep ore in the leases to the south of The Zinc Corporation which were purchased from Barrier South Limited. The Zinc Corporation, Limited and New Broken Hill Consolidated Limited mines are both managed by Conzinc Riotinto of Australia Limited, which wholly owns The Zinc Corporation, Limited and has a 32% interest in New Broken Hill Consolidated Limited. * Throughout this paper tons referred to are long tons of 2,240 lbs.

Jan 1, 1970

By H. W. Rogers

The first steam shovels used in this country were built by the Otis Company, of Boston, about 50 years ago, but as they were of very crude construction and rather unsuccessful only a few were built. For possibly 10 years prior to 1884 successful steam shovels were made and used on certain classes of work, but it was not until that time that they were manufactured in quantities and began to play an important part in all classes of excavation. From that time up to the present day there have been gradual but continuous improvements on the original shovel, not only in the mechanical construction, but also in the design of boilers and engines which are best adapted to this class of service. In proposing a change from steam to electric operation we have to deal with a steam equipment which has not only proved its worth but has probably reached its highest stage of development and efficiency. That the electric shovel is a possibility cannot be denied, as at present from 12 to 18 shovels are in operation in this country. These shovels may be divided into three classes: the friction electric, which is operated by a single constant-speed motor with friction clutches; the three or four motor direct-current equipment, and the three or four motor alternating-current equipment. It is the second and third classes that I wish to deal with, as the first class does not compare favorably with the steam shovel so far as speed is concerned, although it may be operated as cheaply. There is probably no other class of machinery that presents a duty cycle as severe as that of the shovel, which is very short, varying from 7 to 12 sec. on the hoist, from 7 to 12 sec. on the thrust, and from 10 to 18 sec. on the swing, making a complete cycle in from 17 to 30 sec., and the motor to meet these requirements must have a sufficiently low armature inertia to permit of rapid acceleration and quick reversals under small power. It should also be a motor of rugged design, as it must be subjected to severe overloads and shocks and frequent reversals. This is especially true of the hoist motors and, to a lesser degree, of the swing motor; the thrust motor being practically stalled during the digging operation, although it may revolve or overhaul, according to conditions, and is operated at full speed only after the hoisting operation is completed.

Jan 1, 1915

By J. D. Livingston

THE hardening of alloys by the precipitation of a second phase has long been an important technological process. One approach towards improving our understanding of this phenomenon has been a correlation of mechanical properties with such parameters as particle size and inter par ticle spacing.' Several reviews of such data have appeared.2-4 A major limitation to this approach has been that for most alloys the particles are submicroscopic in size in the range of optimum hardness. Precipitation-hardening copper-cobalt alloys, therefore, are of interest, because the ferromagnetic properties of the precipitate make it possible to measure particle size, number of particles, and so forth, from magnetic measurements alone.5, 6 Livingston and Becker7 have applied these techniques to a study of precipitation hardening in an alloy of 2 pct Co in copper. It was found from magnetic measurements that nearly all the cobalt to be precipitated was out of solution within a few minutes at 600°C, whereas the peak in tensile yield stress was not attained until a much longer aging time. Thus during the first 1000 min of aging time at 600°C the Yield stress was increasing while a constant volume of precipitate was coarsening, i.e., while particle size and interparticle spacing were increasing. Peak yield stress was reached at an average particle radius of about 70A. The earlier study was confined to one alloy at one aging temperature, so that the volume fraction of precipitate was fixed and, therefore, particle size and interparticle spacing could not be independently varied. The work has now been extended to several alloys of varying percentage cobalt in copper, and to aging at various temperatures. It was hoped that such a study might clarify the relative importance to precipitation hardening of particle size, interparticle spacing, and volume fraction of precipitate. EXPERIMENT The alloys studied contained 0.7, 1.3, 2.0, and 3.2 wt pct Co in copper. They were prepared from electrolytic copper and electrolytic cobalt in a vacuum -induction furnace, worked to 3/8-in.-diam rod, and machined into tensile specimens with a gage section 1 1/2 in. long and 3/16 in. in diam. All specimens were given a solution treatment consisting of a 30-min anneal in nitrogen at 1000° C followed by a quench into iced brine. Subsequent aging was done in nitrogen and followed by a water quench. Average grain diameter was approximately 0.01 in.

Jan 1, 1960

By Hugh Marriott

MANGANESE and its ores have been recently dealt with in comprehensive papers to the Iron and Steel Institute by Sir Robert Hadfield, and in a series of papers read before the A. I. M. E. at the Cleveland meeting. I have taken the opportunity of adding this mineral, to my records of the "World's Mineral Production and Consumption," as recently pub-lished in the Economist ?. The tables bring out the following significant facts: The world's production has increased by 10 per cent since 1913. India's production has increased by 3 per cent since 1.913, the supply from this source now being 31.1 per cent of the world's output. Russia is now producing at about half its pre-war rate and yields 19 per cent of the world's production. West Africa supplies 12.5 per cent of the world's total output. Brazil's present output is 2.5 times that of the pre-war period, being 11.3 per cent of the world's output.

Jan 9, 1927

By G. H. Gilman

The campaign now being waged to improve the quality of the rock-drill bit is the natural outcome of the scientific development of the drilling machine during the past twenty years. In this development there has been a great increase of power output per unit of weight of machine. But the development of the rock drilling engine has been seriously hampered by the need of a drill bit that would withstand, without undue wear or breakage, the heavy rapid blows of the improved pneumatic hammer. This problem was difficult to solve because an increase in the weight or section of the drill-steel bar resulted in decreased cutting speed and higher operating costs of both the bit and its actuating engine. The solution, therefore, is to make the bit of steel that, for a given weight or section, will be more shock and wear resisting. In the attempt to meet these requirements, many of the drill-steel manufacturers alloyed the steel with vanadium, chromium, nickel, etc. on the assumption that such elements would improve the rock-drill bit but due to the sensitiveness of such steel to forging and heat treatment, higher cost, and lack of continuity of effort on the part of the steel manufacturer to perfect this product, but little progress has been made in the adoption of these steels and the so-called straight carbon steel is today recognized as the standard material for all rock-drill bits. The first step in improving drill-bit efficiency is the determination of the cause of breakage and wear. For this reason, in 1917, certain rock-drill manufacturers, in conjunction with some of the larger mining interests, conducted exhaustive tests in both the field and the laboratory. It was then found that, in the main, the cause of drill-steel troubles is not attributable to the duty imposed upon it by the normal operation of the drilling engine, but to improper forging and heat treating of the bit and to the subjecting of the bit to improper use. The fact that it is possible to minimize the troubles attributable to improper heat treatment of the drill steel has been demonstrated and it is recommended that such knowledge be made available by teaching those interested in the production, use, and up-keep of drill bits the principles on which success in this direction may be achieved. Unfortunately, but little attention has been given to the education of the drill smith. But a few years ago, the instructions given the man

Jan 1, 1922

By Susan B. Leiter

That oxygen in copper has been a source of trouble is well known and that that trouble has been real in the commercial world has been shown by Fuller.' Moore and Beckinsale's paper2 at the annual meeting of the Institute of Metals in London, 1921, called forth considerable discussion. Prof. T. Turner expressed the opinion that it was not necessary to have cuprous oxide present in sufficiently large quantities to form either globules or seams in order to produce a bad effect if the copper were gassed. This brought up the question of the solubility of oxygen in copper and the authors desired to know whether there were any evidence of such solubility, because their experience was that whenever there was any oxygen in copper it was present in the form of free cuprous oxide, which was practically always in the form of globules. Later Hanson, Marryat, and Ford3 stated that 'for all practical purposes cuprous oxide may be regarded as insoluble in copper. However, they found particles of the eutectic to coalesce to form much larger particles. They suggested that cuprous oxide was actually slightly soluble in copper, for it was difficult otherwise to understand the mechanism by which such coalescence could take place. W. E. Ruder, in discussing that paper, said that his experiments had led him to conclude that a slight solubility must exist to allow for the ready coalescence of oxide particles, but conductivity tests convinced him that that solubility was very small. Nevertheless, there appeared to be a state of admixture of oxide with copper that closely resembled a colloidal mixture which at first was mistaken for a true solution and which usually existed at the grain boundaries. The purpose of this paper is to give the results of some investigations, begun in September, 1921, which would seem to show that although cuprous oxide may be only slightly soluble: (1) this solubility is sufficient

Jan 1, 1926

By Enid C. Plante, K. L. Sutherland

PRACTICAL metallurgists are unanimous in stating that oxidation of mined sulphide ore adversely affects separation of the constituent minerals under standard conditions in a mill. Frequently, the need for different operating conditions in different plants is attributed either to the extent of oxidation of the mineral surfaces or to soluble oxidation products. The investigations reported in this paper study the effect of oxidation on the primary condition necessary for flotation, namely that of adhesion of an air bubble to the mineral. OXIDATION PRODUCTS FROM SULPHIDE MINERALS Weinig and Carpenter1 state that the relative rates of oxidation of some sulphides are: FeAsS > FeS2 > CuFeS2 > ZnS >PbS > Cu2S tetrahedrite. They also show that oxidation is rapid at first, that sulphides of older geological formations are more stable than those of recent date and that mixtures oxidize more rapidly than any one mineral alone. The more rapid oxidation of mixtures is borne out by results of Gottschalk and Buehler2 who found that PbS, ZnS and CuS oxidized 8 to 20 times faster when pyrite was present. Only the end products of oxidation, sulphate and the metal ion, were identified in these tests. Gaudin3 states that pyrite is difficult to oxidize in the laboratory. Winchell,4 using pyrite containing 99.98 pct FeS2, allowed water to percolate through a bed of this ground material. At the end of three months only a trace of Fe +++and S04- were found. Even after 10 months the quantity of Fe+++ was only 7.8 mg and S04- 25.5 mg per liter of percolating solution. Carmichael5 confirmed Winchell's results on the slow oxidation of pyrite. Usually the oxidation of sulphide surfaces has been conducted in neutral or acid solutions, the acidity being caused by the oxidation products. Eliseev and Nagirnyak6 found that treatment of pure pyrite with lime gave only thiosulphate, but considered it possible that thiosulphate was obtained not by oxidation of the pyrite itself, but by oxidation of calcium disulphide. Mitrofanov7 suggested that thiosulphate was formed by the following process: oxidation of sulphide to give free sulphur (this frequently happens with marcasite), followed by oxidation of the sulphur to sulphur dioxide and reaction of the sulphur dioxide with sulphur to give thiosulphate. Ferric and ferrous hydroxides, or similar compounds, were shown to be present. The Research Staff of the United Verde Copper Co.8 treated ore from the United Verde mine with lime. Thiosulphate, sulphide and polysulphides were stated to be

Jan 1, 1947

By J. M. McKenzie

p-n junctions to detect and measure the energy of single quanta of electromagnetic radiation are described. Useful energy range is from 1 kev to several Meu. Achieved energy resolution varies from 30 pct at 6 keu to 0.12 pct at 2 Mev. Parameters affecting the perfornzance of the present detectors and desired of a new material are discussed. DURING the last 6 years semiconductor detectors have been used extensively to detect and to measure the energy of charged particles. These detectors are made from reverse-biased p-n junctions. Electron-hole pairs are created in the depletion region of the junction, Fig. 1. The carriers are separated by the electric field and give rise to a pulse of current in the external circuit. This current pulse is proportional to the number of carriers and hence to the energy deposited in the depletion region by the ionizing event. More recently the discovery of the lithium drift process, first in silicon' and then in germanium,2 has extended the use of semiconductor detectors to the detection and the energy measurement of single quanta of electromagnetic radiation. These detectors now have a useful measurement range extending from the soft X-ray region, about 6 kev, to the hard y ray region of several Mev. Surface-contoured avalanche diodes can be used to extend this range down to 1 kev. There is however a desire for other materials suited to higher operating temperatures, and particularly for those materials with high y ray stopping power. The most important advantage of the semiconductor detector over gaseous and scintillation detectors is the average energy, E, required to form an electron-hole pair. Fig. 2 shows the quantum efficiency for germanium as a function of photon energy.3 At the band-gap energy, 0.7 ev, an electron-hole pair is generated by the absorption of the photon and the quantum efficiency is unity. The quantum efficiency remains constant at unity between 0.7 and 3 ev. It then increases as additional electron-hole pairs are formed by secondary ionization of the initial photoelectron. The rate of increase is about 1 electron-hole pair per 3 ev (actually 2.9 ev for germanium and 3.6 ev for silicon). This figure is constant irrespective of the particle type (electron, proton, a particle, heavy ion) and energy. In an ionizing event an incident particle of energy E will create N = E/e ion pairs. The statistical fluctuation in this number is proportional to &. The fractional energy resolution is therefore proportional

Jan 1, 1968

By Gustav S. Preller, A. M. Gaudin

For the past 20 years it has been generally accepted that the flotation process is made possible as a result of the action of certain chemical substances on the surface of the mineral particles. In fact, it has been suspected, but never demonstrated, that this action produces on the mineral particles a layer of chemosorbed ions that is of the order of magnitude of a monolayer; that is, of a layer one ion deep. It is the purpose of this paper to present certain initial results on the relationship of mineral surface area and collecting-agent requirements. Up to about eight years ago, all considerations involving the surface area of finely divided solids were limited by the fact that there was not in existence any general method for determining the extent of such surfaces, especially on mixed materials. However, the method invented by Brunauer, Emmett and Teller1 has changed the situation by providing a general procedure (BET method). This procedure has already been applied to mineral dressing problems2-4 and has now received confirmation as to accuracy by an entirely independent method.' This confirmatory or "absolute" method uses a very tedious calorimetric procedure and is not suitable for routine determinations; but still another method6 now is available to relate the results obtained by the BET method to those obtained calorimetrically. This Harkins-Jura method has the advantage that it does not require an assumption concerning the parking area per molecule of adsorbate (nitrogen). In the original work by Brunauer, Emmett and Teller the parking area per nitrogen molecule was calculated to be either 13.8?2 or 16.2?2. according as the nitrogen was assumed to be closely packed as in crystalline nitrogen or loosely packed as in liquid nitrogen. The initial experimental work from this laboratory2 supported the 13.8?2. figure; that work was done on glass spheres of measurable geometric area and it was based on the assumption that the glass spheres were truly smooth; any deviation from this assumption would, of course, have made the calculated parking area higher. Harkins and Jura's results suggest that the parking area is not the same for all substrata; furthermore, the average of the 87 measurements cited indicates a "preferred" parking area for the nitrogen molecule of about 15.2?2., with extreme values at 13.7 and 16.9?2. The average of 21 determinations made in connection with this study7 gave a parking area of 14.3 ± 0.27?². In view of this result and of the Harkins and Jura

Jan 1, 1947