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By J. W. Cuthbertson, R. A. Cresswell

THE constitution of Cu-rich alloys with 1.5 to 13.5 pct Sn and 0.25 to 3.0 pct Be and the precipitation-hardening characteristics of alloys with 1.5 to 13.5 pct Sn and 0.25 to 1.0 pct Be have been examined. The hardness and tensile strength of the alloys examined increase markedly after solution treatment at 700°C followed by heat treatment at temperatures between 200" and 450°C. By a combination of cold work and heat treatment, hardness values similar to those exhibited by commercial Be-Cu alloys containing 2.25 pct Be can be obtained with ternary alloys containing 9 pct Sn and 0.75 pct Be and containing 10 pct Sn and 0.5 pct Be. Marked hardening effects occur with alloys containing even less beryllium. By heat treatment alone, a hardness value of 310 diamond pyramid hardness can be obtained from an alloy containing 10 pct Sn and 0.75 pct Be. Preliminary tensile tests have shown that an ultimate tensile strength of 110,000 psi with an elongation of 23 pct is obtainable by precipitation hardening an alloy with 8 pct Sn and 0.75 pct Be. The precipitation-hardening process has been followed microscopically for certain alloys and the inference is that, while the initial hardening effect is probably explained by the precipitation of the ß phase of the Cu-Be system, further hardening, proceeding at a much slower rate, also occurs, apparently as a result of precipitation of phases of the Cu-Sn system, particularly precipitation of the 6 phase at temperatures below 350". The presence of the e phase of the Cu-Sn system in certain alloys at temperatures below 350°C has been confirmed. Tin-bronzes are widely used in engineering applications where a combination of high strength and good resistance to corrosion is wanted. The maximum strength is induced in these alloys by cold working, and it would be an advantage for many purposes if high strength could be achieved alternatively by an age-hardening process. While Cu-Sn alloys have a good fatigue resistance they can be surpassed in this respect by Cu-Be, but the use of the latter alloy is limited by its high cost. If, by adding beryllium to tin-bronze, the properties of the respective binary alloys could to some extent be combined, a most attractive alloy should result. As pointed out by Raynor,¹ beryllium is on the borderline of the zone of favorable size factors for copper, and the solid solubility of beryllium in copper is consequently much more restricted than if the size factor were strongly favorable. The size factor is sufficiently favorable, however, to permit an increase in solid solubility with rise in temperature, and there is thus a composition range in which CU- Be alloys are susceptible to hardening by precipitation heat treatment. Although the a phase of the Cu-Sn system is similarly susceptible to precipitation treatment, the time necessary to establish equilibrium in commercial alloys of this type is usually so great that age hardening becomes impracticable. The addition of beryllium to Cu-Sn alloys would appear to offer a means of conferring on the latter useful age-hardening properties. Masing and Dahl² and others have, in fact, shown that the addition of beryllium to Cu-Sn a solid solutions renders these alloys susceptible to precipitation hardening and after such hardening confers on them an encouraging improvement in physical properties. If this improvement could be achieved by the addition of substantially smaller amounts of beryllium than are customarily found in binary Cu-Be alloys, the ternary alloys should possess economic advantages which might make them more attractive than the binary alloy for some applications. Binary Systems Copper-Tin: The constitution of these alloys is now reasonably well known and is summarized in the equilibrium diagram published by Raynor.³ The following observations, due to Raynor,¹ on the structure of those phases of the Cu-Sn system that are likely to be found in the ternary alloy system will facilitate the subsequent discussion on the examination of that system. The ß phase is an electron compound at the electron-atom ratio 3:2 and has a body-centered cubic crystal structure. This phase is stable only down to 586°C, at which temperature it decomposes eutectoidally into the a and y phases. The y phase has a structure that is also based on the cubic system. This phase is stable down to 520°C, at which temperature it decomposes eutectoidally into the a and d phases. The d phase is an electron compound (Cu³¹Sn8) which has a crystal structure analogous to that of 7 brass. This phase is stable from 590" to 350°C; on prolonged annealing at the latter temperature it breaks down into a mixture of the a and E phases. The e phase is an electron compound (Cu³Sn) having the electron-atom ratio 7:4. Its structure may be regarded as a superlattice based on the close-packed hexagonal system. This phase is stable from 676°C to room temperature. The primary solid solubility of tin in copper increases to a maximum of 15.8 pct as the temperature falls from that of the peritectic reaction to 586°C. The solid solubility remains constant from 586" to 520°C. At lower temperatures the solubility decreases progressively. Below 350°C the fall in solubility is pronounced and is associated with the precipitation of the e phase. This precipitation is very sluggish and does not normally occur under service conditions. Copper-Beryllium: The Cu-Be system has been investigated by Borchers' and others. Raynor5 summarized the present state of information on it.

Jan 1, 1952

Edward Halse, Puerto Berrio, Colombia, So. Amer. (communication to the Secretary*): The Taviche mining-district is generally known as " Taviches," being named from the two villages called respectively San Pedro Taviche and San Jero-nimo Taviche. It is so called by Jose G. Aguilera, in his paper on the Geographical and Geological Distribution of the Mineral Deposits of Mexico;' and in published reports by the late N. W. Rudston Read ;2 and by Juan de Dios Villarello3 The region is a large one, measuring about 24 miles long and 7 miles wide. Dr. Chance, in his interesting paper, tells us very little about the nature of the ores, and, with regard to the geology of the district, contents himself with observing that " the country, as a rule, consists of eruptive rocks; and no evidences of sedimentary strata appear near the mines." The latter part of the above statement is hardly correct. The formation is mainly

Jan 1, 1906

By J. Wes Blakely

In the mid-1950s the importation of foreign oil began to encroach on traditional coal markets such as railroad steam engine fuel, home heating, electrical generating plants, and others. Many of the smaller coal companies with substantial coal reserves subsequently found that they could no longer continue to mine economically and merged in various ways with larger established coal mining concerns. As the seats of top management and authority were removed from the mines to large metro centers such as New York, Pittsburgh, and Chicago, the more personal relationship between mine workers and management that had existed prior to these mergers began to be eroded and replaced by "remote control" mine management. The disappearance of this face-to-face contact of upper management with mine employees tended to create barriers blocking simple effective communications between coal mine management and labor. In examining the human relations aspects of the coal mining industry, many elements, some concerned with work and others not, have depicted the industry as almost devoid of good applied human relations and communications practices, and, in truth, little has been done to alleviate this situation. In past years many coal producing companies, with the exception of one of the largest, failed to adequately define the term "communications ' as it applied to their companies, whether internally or externally. Internally many still place heavy reliance on written communications such as company-wide newspapers to do the job of communicating with their employees which they feel ought to be done. One company puts an inordinate belief in its company "newspaper" as the primary instrument to stimulate its work force into increasing productivity. Primarily aimed at the workers, who greatly outnumber the salaried administrative people, this publication mainly documents the fortunes of individual mine workers who are felt to have achieved something with their lives, e.g., a mine worker who had managed to put four children through universities, the

Jan 1, 1981

By J. Wes Blakely

In the mid-1950s the importation of foreign oil began to encroach on traditional coal markets such as railroad steam engine fuel, home heating, electrical generating plants, and others. Many of the smaller coal companies with substantial coal reserves subsequently found that they could no longer continue to mine economically and merged in various ways with larger established coal mining concerns. As the seats of top management and authority were removed from the mines to large metro centers such as New York, Pittsburgh, and Chicago, the more personal relationship between mine workers and management that had existed prior to these mergers began to be eroded and replaced by 'remote control mine management. The disappearance of this face-to-face contact of upper management with mine employees tended to create barriers blocking simple effective communications between coal mine management and labor. In examining the human relations aspects of the coal mining industry, many elements, some concerned with work and others not, have depicted the industry as almost devoid of good applied human relations and communications practices, and, in truth, little has been done to alleviate this situation . In past years many coal producing companies, with the exception of one of the largest, failed to adequately define the term “communications” as it applied to their companies, whether internally or externally. Internally many still place heavy reliance on written communications such as company-wide newspapers to do the job of communicating with their employees which they feel ought to be done. One company puts an inordinate belief in its company "newspaper" as the primary instrument to stimulate its work force into increasing productivity. Primarily aimed at the workers, who greatly outnumber the salaried administrative people, this publication mainly documents the fortunes of individual mine workers who are felt to have achieved something with their lives, e.g., a mine worker who had managed to put four children through universities, the

Jan 1, 1981

By Edward Hart

Having occasion several years since to make ferrous sulphide, I attempted to do so by fusing a mixture of coal-brasses (FeS2) and dried ferrous sulphate. A very nice-looking sulphide was obtained; but on attempting to use it, I found that it was insoluble in dilute hydrochloric and sulphuric acids. The subject was laid aside at the time, but has since been taken up; and some curious results have been obtained, which I believe are new. I have made the material by all of the following methods and the product seems to be pretty much the same, however obtained : (a) FeS2 heated to fusion. (b) FeS2, and FeSO4 heated to fusion. (c) FeS2, and Fe2O3 heated to fusion. (d) FeS2, and FeCO3, heated to fusion. (e) FeSO4, and S heated to fusion. (f) Fe2O3 and S heated to fusion. Obtained in this way the mass after cooling breaks easily and is very nearly black in color. On the upper surface of the fused cake the color is generally reddish or brownish black. When this is treated with dilute sulphuric or hydrochloric acid, it refuses to dissolve; but if touched with the point of a file or any piece of wrought-iron, solution of the sulphide takes place with the evolution of sulphuretted hydrogen. If the iron is removed the evolution of gas ceases, and only begins again when the iron is again brought into contact with the fused sulphide. The same thing happens if the iron is connected with the sulphide by means of a platinum wire. But in all the samples of fused sulphide I have obtained, if the iron is allowed to remain in contact with the sulphide for a certain time (which varies with different samples of the fused sulphide) and is then removed, solution of the sulphide continues. Zinc, tin, and magnesium have the same effect as iron. In hydrochloric acid aluminium acts more rapidly even than iron. In sulphuric acid it has little if any effect. Copper in HCl, when brought in contact with the fused sulphide, dissolves quietly without evolution of gas. The solution of copper chloride formed becomes,as it flows down over the sulphide, converted into copper sulphide.

Jan 1, 1887

By G. L. F. Powell, L. M. Hogan

Samples of silver and Ag-O alloy, 0.12 wl pet, have been undercooled to a maximum of 250°C by melting in a slag of commercial soda-lime glass. Grain refinement occurred in undercooled silver samples whet the degree of undercooling exceeded a critical value which varied from 153° to 175°C, depending on the melting conditions. In under cooled Ag-0 samples, the grain structure was fine and equiaxed at degrees of undercooling larger than 50oC. The grain refinement in both silver and Ag-0 alloy samples was the result of recrystallization during or immediately following freezing. When the oxygen content was increased, the degree of undercooling at which recrys-tallization occurred decreased. INVESTIGATIONS of the grain structure of strongly undercooled melts of nickel and cobalt1-4 have revealed a change in room-temperature grain size as a function of undercooling. At large degrees of undercooling the grain size was fine, but samples undercooled by smaller amounts exhibited a much coarser grain structure. The reported undercooling at which this grain size transition occurred in nickel varied between 140° and 175'C undercooling. Walker1 attributed this grain size change to an enhanced nuclea-tion rate. Cavitation at a melt-solid interface was con sidered to generate a high-pressure wave. This, in turn, produced nucleation in the liquid as a result of an elevated freezing point increasing the undercooling in the melt. A similar grain size transition was observed in undercooled silver between 133° and 153°C undercooling.5 The grain size transition in undercooled silver was found to be dependent on oxygen con tent. More recently, the authors5 showed that a simila transition occurred in an undercooled Cu-O alloy, =0.08 pet, but not in oxygen-free copper undercooled more than 200°C. The transition was attributed to recrystallization of the solidified grains after freezing; the greater the degree of undercooling, the more complete the recrystallization. In view of this effect of oxygen content on grain size transition, additional undercooling experiments were carried out with silver. This paper reports the results of these studies on silver and Ag-O alloys undercooled more than 50°C EXPERIMENTAL 1) Materials. The silver was supplied by Matthey Garrett Pty., Ltd., as fine granulate, silver mini' mum 99.9 pet, and contained the trace impurities shown in Table I. The silver originated from two different refineries resulting in the difference in impurity content shown in Table I. However, the variation in composition had no discernible effect on the undercooling behavior. Samples of the silver were melted in an open-ended, vertical, cylindrical resistance furnace, wound with Kantha] wire, the temperature of which was adjusted by means of a variable transformer, 2) Melting and Undercooling. Bulk samples of silver were undercooled by preoxidation and removal of oxidized impurities in a glass slag, as previously described,5 but special precautions were necessary to allow comparison between oxygen-free silver and a Ag-O alloy. Samples free of oxygen were prepared by first melting weights of granulate varying from 200 to 400 g in clean vitreous silica crucibles in air for approximately 15 min. A sample was frozen and a complete glass cover formed over the top surface by adding granulated glass rod. The sample was then melted and frozen several times until no oxygen bubbles appeared in the glass slag. Subsequent thermal analysis indicated that the oxygen content was negligible. The glass acted as an efficient barrier to absorption of oxygen from the air. Ag-O alloys were prepared by melting silver in silica crucibles with only a partial cover of glass, so that the top surface of the melt was in contact with air. The silver was held molten in air for several hours at 1000° to 1050°C, in the expectation that the oxygen in solution would reach equilibrium with atmospheric oxygen. At T - 1000°C the equilibrium oxygen content was calculated as 0.12 wt pet O, using Sieverts law and Eq, [l], derived by Mizikar et a!.7 from the data of Sieverts and Hagenacker8 for 1 atm of oxygen:

Jan 1, 1970

By E. H. Crabtree, T. G. Chapman, V. W. Winters

This paper traces the experimental and pilot plant work completed by the authors since 1939 including the various methods which have been developed in applying coarse activated carbon to cyanidation. The methods and results developed by the authors in desorbing gold and silver from carbons are given and the possibilities for the re-use of the desorbed carbons are discussed.

Jan 2, 1950

By E. J. Kenaedy, E. H. Robie

THOUGH the fall meeting of the Coal Division was held in the heart of the anthracite section, at Hazleton, Pa., the bituminous industry was well represented also, and the two groups found much common ground for discussion. The' technical part of the program was confined to two sessions in the morning and afternoon of the first day, Friday, Oct. 14, the evening and the following day being reserved for field trips and social functions. The attendance more than met the expectations of those in charge of the arrangements; there seemed no hint of depression here.

Jan 1, 1932

By W. R. Atkins

The Negaunee mine is at the east end of the Marquette Range, in the city of Negaunee, on the Upper Peninsula of Michigan. Iron ore was first discovered on this property in 1883 by diamond drilling. A shaft was sunk soon afterward. and the mine has been worked continuously ever since. The total shipments from 1887 to Jan. I, 1945, were 20,417,587 long tons. The mine was taken over by the present operators, The Cleveland-Cliffs Iron Co., in 1903. During the past four years, 1941 to 1944 inclusive, the property has produced an average of nearly a million tons a year, and has employed more than 400 men. Geology The ore body at the Negaunee mine lies in a large syncline dipping about 15° to the west, the ore coming to ledge at the east end of the trough. The ore occurs in the Negaunee formation of the middle Huronian age, which consists of ferruginous cherts and slates through which intrusions of diorite have occurred. Although these intrusions of diorite have an important bearing on the formation of ore on the range, they are not important in this ore body. Some faulting has taken place along the diorite intrusions, or dikes as they are called locally, and also other faults are known to exist across the formation. These faults do not complicate either mining or development of the ore body. The footwall consists of jasper underlain with Siamo slate; the hanging wall is jasper and ferruginous chert. Fig. I shows a typical cross section of the ore body. The ore body is quite regular, about 120 ft. thick in the middle of the syncline and thinning out toward both sides of the fold. Its width ranges from 400 to 800 ft., the average being about 600 ft. The footwall is uniform in dip and strike, but irregular lenses of ore often extend 50 to 75 ft. up into the overlying jasper capping. Method of Mining Ore in the Negaunee mine has been mined by top slicing almost exclusively since 1903 This method of mining has been employed because it is especially adapted to the size and shape of the Negaunee ore body, and to the soft, sticky nature of the ore as well as the generally soft, broken jasper capping. Development The ore body has been developed from six different levels with intervals varying between loo and 125 ft. On the average, the dip causes a horizontal offset of 350 ft. per level, and under these circumstances it has been necessary to carry on considerable development in rock to make available the ore lying along the north and east footwalls. This is not as disadvantageous as might be supposed, because of reduced pressure and the accompanying saving in maintenance costs. The development of the

Jan 1, 1949

By Jei Y. Choi, Paul G. Shewmon

The surface self-diffusion coefficient of copper (D,) has been measured between 847° and 1069 "C for six different orientations. These were the(111), (110, (100, and three higher index surfaces. The activation energy for Ds (designated Q s) was found to be about 49 kcal per mol for all six surfaces, and Do about 2 x 104 sq cm per sec. At any temperature Ds varied by no more than a factor of three over these orientations. It is shown that, if the free energy of a surface atom is uniquely determined by its number of nearest neighbors, it follows from the Principle of microscopic reversibility that Qs should have the same value for all surface orientations, and Ds should vary little with orientation. This model also suggests that for clean fee metals Qs ~ 2/3 AH, (heat of vaporization). This is true for copper. ALTHOUGH it has been appreciated for several decades that atoms can diffuse more rapidly on a surface than through the bulk of a crystal, it has only been in the last few years that reliable values of the surface self-diffusion coefficient (Ds) have become available. Tracer studies of Ds had been attempted prior to this period, but when a tracer is placed on a surface, an ever increasing fraction of it is drained off into the lattice. The correction for this loss involves a very difficult, and as yet unperformed calculation. Those who have worked with tracers have not corrected for this loss.1, 2 Thus their results indicate that Ds is greater than the self-diffusion coefficient in the lattice (Dl), but it has not been established that they give quantitative data on Ds. A procedure which avoids the problem of tracer loss is to study the rate of mass-transfer under the effect of surface tension. If the surface asperity being studied is very small, the mass transfer occurs entirely by surface diffusion. The kinetics at which a grain boundary groove forms on an initially plane surface is a well-studied case of this type. The smoothing of a slight scratch in an otherwise flat surface is another procedure that has been studied. If these grooves are up to 20 to 30 µ in width, the dominant mechanism for mass transfer is surface diffusion (at least in the case of metals with low vapor pressures), and the widths can easily be measured with an interference microscope. Of these two, mass-transfer techniques only in the case of grain boundary grooving has a rigorous mathematical treatment been given. This was done by Mullins.3,4 His analysis predicted that in the case of copper in an atmosphere of an inert gas, surface diffusion should be the dominant transport mechanism. This analysis gave an equation for the groove profile and predicted that the width of the groove would increase as (time)1/4. Mullins and Shewmon showed that both of these predictions agreed with experiments.5 Thus the validity of the values of Ds given by this procedure seems to be well established. Gjostein has used copper bicrystals and the grain boundary grooving technique to determine Ds and the activation energy for surface selfdiffusion (9,) in the [001] direction on surfaces ranging between the (100) and (110) planes.= He reported that Qs = 41 kcal per mole and Do = 6.5 x 102 sq cm per sec for all orientations studied. Since the results did not change with the dew-point of the dry hydrogen atmosphere or the type of refractory tube used, he concluded that the surfaces were clean, or at least that the results were not influenced by any impurities chemisorbed from the atmosphere. The work reported here reproduces and extends Gjostein's study in that D s and Q s were determined for copper over a wider range of orientations. To study the effects of impurities, two purities of copper were used as well as cathodic etching to remove any possible electropolishing film. Gjostein postulated that the diffusing atoms on a surface near a low index plane are the few atoms which are adsorbed on the smooth region between ledges or steps in the surface. A more rigorous derivation of the equation relating Ds to the concentration and jump frequency of these adsorbed atoms is given here. Using this treatment, our empirical observation that Q s and D s are essentially the same for all surface orientations can be shown to follow from the assumption that the free energy of a surface atom is uniquely determined by its number of nearest neighbors. The studies of D s using the scratch technique have been carried out by Blakely and Mukura on nickel,' and by Geguzin and Oveharenko on copper. The latter study using copper gives values of D s roughly

Jan 1, 1962

By Charles Barron McIntosh

BECAUSE of the well known relationship between a low atmospheric pressure and increased amounts of methane in coal mines, attempts have been made to find associations between low pressures and coal mine explosions. An early study by F. Able disclaimed this relationship. He found that half the explosions took place during a rising barometer when methane content in a ventilated mine would be decreasing. A more recent study by C. L. Hosler revealed a close relationship between anthracite mine explosions and low atmospheric pressure. Bituminous coal mine explosions, where coal dust as well as methane can be exploded, were found to be more closely associated with a rising pressure following a low barometer than with the falling or low pressure period. Emphasis on rising pressure periods led the present writer to investigate the effect of the post-frontal or post-cyclonic weather on coal dust. Dust samples of uniform mesh were exposed in an Illinois mine. Analyses of the dust taken before and after the passage of cold fronts and their accompanying low pressure troughs indicated an increase in coal dust moisture prior to a frontal passage and loss of moisture from coal dust following the time of lowest pressure. Thus the warm, moist air (Tropical Maritime)found over the mine during the falling pressure period was cooled in the mine and moisture added to the dust. Cold, dry air (Polar Continental) accompanies' the rising barometer following a frontal passage and the air, being heated in the mine, subtracted moisture from the coal dust. A similar but seasonal phenomenon occurs from summer to winter, the dry dust period in winter being recognized as the more dangerous period. The question arose as to how the short period or daily changes in moisture content of coal dust compared with the seasonal changes. Temperature and humidity statistics were obtained from the surface and mine entry for the months of July andJanuary. The gain and Ioss of mdisture. to the hine were calculated for these two morths as were the gain and loss associated with mT (Tropical Maritime) and cP (Polar Continental),air masses as observed in the investigatiynsl. Temperature and humidity statistics tigatiynsl.taken from [the two air masses represent atimp span of only 34 hr. The redultant changes as presented in the accompanying table verify the supposition that short period changes in coal mine moisture may attain the magnitude of recognized dangerous seasonal changes. Statistics for the mT air and the July air have been placed next each other and those for cP and January air together for easier comparison of the two warm and two cold air units. The last column readily reveals that there is little difference in the amount of moisture being deposited in the mine from the mT air and the July average air. A similar conclusion is reached regarding the loss of moisture from the mine when it is affected by the cP air mass or the average January air. This comparison would be particularly true from late fall through early spring when strongly contrasting air mass changes are a common occurrence. The invasion of cold, dry air following a cold front thus produces a drying period that could well receive more attention as a danger period for coal dust explosions. This drying period would compare with Able's explosion period during a rising barometer. The drying period is also perltinent in Hosler's findings, where it was noted that the average explosion day (for bituminous coal mines) followed the day of lowest atmospheric pressure. Moreover, the author's investigation of some 365 coal mine explosions also indicates that this post-frontal period should receive special attention in connection with mine safety measures. Two comparatively recent major explosions in Illinois (Centralia and West Frapkfort) occurred after the mines had been under the influence of cold-dry air for 25 to 30 hr-following a deep low pressure passage. Dust was the important explosive element in both these explosions.

Jan 1, 1958

By K. E. Merklin, F. D. DeVaney

Students of the modern blast furnace seem unanimously agreed that they are observing a major revolution in practice. Rather than changing construction and operation of the furnaces, most of the great advances now under way deal with the raw materials feed. Prepared burdens have gained increasing significance during recent years. Sizing, concentration, and agglomeration have all become bywords to greater production and reduced operating costs. The remarkable results with test furnaces in which self-fluxing sinter makes up part or all of the ore charge foretell even greater improvements. Other investigators have reported many of these advances, and it is enough to predict here that more and more operators will turn to the use of self-fluxing sinter. This change in the operator's requirements comes at a time when one major section of the iron ore regions, the Mesabi, is finding it difficult to maintain, let alone improve, the tonnage and grade of its shipping product. To meet the new demands with respect to quality and structure, the iron ore producers of this area turned first to screening, crushing, and concentrating intermediate ores. More recently, declining higher-grade reserves have led to multimillion-dollar investments in plants to treat the low-grade magnetic taconites.1,2 To yield a suitable concentrate, the Minnesota magnetic taconites must be ground to about 80 pct through 325 mesh, too fine to be charged directly to the blast furnace or to make acceptable feed to the sinter strand. The pelletizing process was developed to meet the structure requirements laid down by the blast furnace operators. Taconite producers are succeeding in their efforts just when the attention of the blast furnace man is being drawn to the great advantages of self-fluxing sinter.3 The taconite operators are ready with an alternative. Why not produce sclf-flus ing pellets? At the Hibbing laboratory of Pickands Mather & Co., where much of the early work was done that led to Erie Mining's 7.5-million ton taconite plant at Hoyt Lakes, Minn., preliminary investigations have been made. Present Taconite Operations: Details of this work will be clarified by a brief review of agglomeration practice in commercial taconite plants. After the taconite has been ground to about 80 pct through 325 mesh to liberate the magnetite grains, the ore is concentrated on magnetic separators, thickened, and filtered to about 10 pct moisture. This structure and moisture fit in closely with the conditions required to produce good balls. It has been found with taconite concentrates that at least 60 pct of the feed to the balling drum must be —325 mesh to obtain good ball compaction within reasonable time. The filter cake moisture usually permits a slight water addition at the balling drum, and further balling control is obtained by adding about 12 Ib of bentonite and 1 1/2 lb of soda ash per dry short ton of concentrate. These additives, and occasionally small quantities of fine coal, are thoroughly mixed with the filter cake ahead of balling. The swelling property of bentonite improves the green and dry strength of the balls so they can be handled without breakage until they are fired. Soda ash is added for pH control to insure maximum performance of the bentonite, and coal is added for more heat in the firing process when required. The balls, normally 3/8 to l-in. diam, are brought to a temperature of 2350" to 2500°F in the indurating furnace. Heat for this process comes from three sources: 1) fuel burned in the combustion chamber, 2) oxidation of the magnetite concentrates to hematite, and 3) coal added to the balls. Pellet strength is developed by grain growth and bridging and, if enough slag-forming constituents are present, by slag bonding. Conversion to hematite is usually close to 100 pct in the fired pellets. Batch Test Procedures: Initial studies of self-fluxing pellets were made by producing balls on a batch basis under carefully controlled laboratory conditions. The l-in. balls, uniform in size, were fired in an oxidizing atmosphere in a standard laboratory muffle furnace. The firing procedure consisted of

Jan 1, 1961

OFFICIAL INSTITUTE REPORTS FOR THE YEAR 1924 Report of the Secretary TO THE BOARD OF DIRECTORS OF THE AMERICAN INSTITUTE OF MINING AND METALLURGICAL ENGINEERS Gentlemen -To a Board of Directors keeping in as close touch with all of the affairs under their care as do the Directors of the A -I M. E, through their monthly meetings, a report of the year's progress is almost superfluous To the members who read the monthly reports of the Board Meetings and of the various activities of the Institute as published each month, a yearly report cannot carry much fresh infor¬mation, yet for both it may be well, as is the custom, to glance over the field of our activities during the past year, and to make an estimate for the year to come The beginning of 1924 was not propitious for the members of the mining profession, though tonnage, production for 1923 had been high and the gross value higher than for several years The margin of profit had been narrow, prospecting and development were limited, and many of our members were unemployed or were working in other industries Toward the end of the year, higher metal prices were in sight Old-time optimism began to grow, and it now brightens the outlook for 1925 An excellent barometer of the prosperity of the industry, and that of the Institute, for they go hand in hand, is the work of our Employment Bureau During the past few months, the percentage of our members for whom openings have been found has been increasing to a marked extent, and many members who had fallen in arrears are placing themselves in good standing The number of new members joining has remained about as it was in 1922 and 1923 The Directors have reason to congratulate themselves on the interest maintained during a dull period, for it is not humanly possible to take a keen interest in the technology of an industry when our hands and minds are not actively engaged in it Now, with a much brighter outlook, we can look forward to a year of increasing prosperity for the industry and for the Institute

Jan 1, 1923

By G. W. Healy, D. C. Hilty

MAJOR considerations in the production of stainless steel are the utilization of stainless steel scrap and the recovery of chromium and other metallic values from the initial furnace charge. The decarburization of a bath of molten chromium steel with gaseous oxygen is usually accompanied by a considerable amount of metallic oxidation. Since recovery of this metal from the slag may require substantial expenditures of time and materials for slag reduction, holding metallic oxidation to the minimum consistent with good operating practice is generally desirable. The important role of temperature in minimizing metallic oxidation is illustrated by Fig. 1, taken from the results of Hilty, Healy, and Crafts,' who established the quantitative relations connecting the amounts of chromium, manganese, and iron oxidized during the oxidizing period of chromium steel heats with the composition of the initial charge and the temperature and composition of the molten bath after oxidation. Higher temperatures after the oxygen blow result in lower metallic oxidation. Although high temperatures are desirable for minimizing metallic oxidation, practical operating considerations limit the degree to which they can be employed. Excessively high temperatures or prolonged operation at moderately high temperatures may cause serious damage to the furnace refractories, so that a compromise must be effected. The problem, therefore, is one of controlling the oxidizing period of chromium steel heats to the shortest time at the highest practical temperature. The most effective tool the steel melter has for attaining high temperature while decarburizing a chromium steel heat is now commonly considered to be the oxygen lance with an adequate supply of gaseous oxygen. In order for the tool to be effective, however, the heat resulting from the oxidation of carbon, chromium, silicon, manganese, and iron in the bath must exceed the heat losses from the furnace. General recognition, of this principle has stimulated a trend toward faster injection of oxygen. For example, Pachaly2 proposed a high temperature at the start of the oxygen blow or a high rate of oxygen input as a means of obtaining the desired carbon level with high chromium recovery. On the other hand, limited understanding of the specific effects of oxygen input rate has hampered development of optimum practices. As part of a broad study of the metallurgy of chromium steel melting, therefore, the Linde Air Products Co. and the Electro Metallurgical Co. at their Metals Research Labo-ratories have evaluated the heat-balance during the oxidizing period in terms of oxygen input rate. The evaluation has then been related to factors such as temperature increase, metallic oxidation, oxygen Utilization, etc., and extrapolated to include a wide

Jan 1, 1958

By Reino Kurppa

The Pyhäsalmi mine of the Outokumpu Co. is located in central Finland (Fig. 1). The deposit was discovered in the autumn of 1958. Construction of the mine was started a year later, and production began on Mar. 1, 1962 (Fig. 2). At present, 750,000 metric tons (mt) of ore and a nearly equal amount of country rock are mined annually. Slightly over one-half million tons of various concentrates are produced per year, of which 85% is pyrite concentrate, 10% zinc concentrate, and 5% copper concentrate. The pyrite concentrate is treated in the company's pyrite smelter,1 yielding elementary sulfur, sulfur dioxide, and iron-oxide cinders, as well as electric power. Thus all the sulfides, totaling 75% of the ore, are of economic value. This is a fact which had a bearing on the development of the stoping method. GEOLOGY The ore is of breccia type containing fragments of country rock of varying sizes. It is coarse-grained pyrite ore containing sphalerite and chalcopyrite as a matrix between the pyrite grains. Only small amounts of pyrrhotite have been observed in certain marginal parts of the orebody. The ore averages 0.85% Cu, 3.0% Zn, 38% S, and 33% Fe. The orebody extends from the surface down to at least 500 m, which is the lowest level from which present information has been obtained. Horizontally, the orebody is 650 m long. It is 75 m wide at its central part, narrowing towards the ends (Fig. 3). At a depth of 115 m, the orebody suddenly narrows down to a thickness of 25 to 30 m. In the northern part of the orebody, the dip is 50° to 70°, but elsewhere it is almost vertical (Fig. 3).

Jan 1, 1969

By Theodore B. Comstock

Prior to 1885, the College of Engineering in the State University was under the care of such professors and instructors as were required for efficient work in the Schools of Mechanical Engineering, Civil Engineering and Architecture. The development of the coal and other important mineral resources of the commonwealth led gradually to a considerable demand for instruction in the branches more closely related to mining engineering, and a very creditable effort was made to meet this want by arranging special work in the field and in the laboratory for students of mining and metallurgy. This plan, while it gave much added labor to several already heavily taxed professors, caused such students to become something like undesirable excrescences upon the regular courses. But its partial success gave abundant proof of the real necessity for a complete course in Mining Engineering. The famous Braidwood disaster emphasized anew the importance of thorough training on the part of those to whom the responsible duties of mining must be confided. As is well known, a very excellent beginning was made some years ago by the Legislature in the matters of inspection and regulation by the enactment of a reasonably effective law, which was afterward much improved.* About the date of the last Acts, it was deemed wise to enlarge the facilities at the University arid to furnish adequate training for the increasing number of applicants. Accordingly, in August, 1885, the writer was requested to assume the duties of the professorship, it being the expressed desire of the Board of Trustees and the Regent of the University that a School of Mining Engineering should be founded which might eventually come to rank with the other engineering schools already flourishing here. It is the main object of this paper to state plainly the steps thus

Jan 1, 1887

By W. A. Barrows, Carl Zapffe

Jan 1, 1913

By John L. Schroder

INTRODUCTION-SYSTEM DESIGN The design of an underground mining operation requires the integration of transportation, ventilation, ground control, and mining methods to form a system which provides the highest possible degree of safety for mine personnel; the lowest cost per ton of product ; the most suitable quality of final product, considering its intended use; the maximum percent recovery of the reserve, consistent with cost and subsidence rights ; and the minimum disturbance to the environment. The selection of mining method is generally the controlling factor in overall mine design. Some important considerations in the choice of mining method are: 1) Depth of cover over the coal seam. 2) Strength and thickness of rock strata in the overburden. 3) Characteristics of the coal seam, including thickness, nature and location of impurities or irregularities, dip (or pitch), nature and direction of cleavage, hardness, and structural strength of the coal. 4) Nature and strength of the immediate roof (top) over the coal seam. 5) Nature and strength of the floor (bottom) under the coal seam. 6) Presence and amount of explosive gas (methane). 7) Quantity of water likely to be encountered in the mining operation. 8) Any previous or concurrent mining in seams above or below the coal seam to be worked. 9) Any restrictions on recovery required for subsidence control. Topographical and geological reports and maps of the area are usually available from state and federal geological surveys, and contain excellent general information to guide more detailed investigations. Exploratory drilling, and the plotting and analysis of the information thus obtained, are essential steps in any development program, and aid in the correlation and evaluation of information gained by other means. Other good sources of information are operating coal mines that may be in the same

Jan 1, 1981

By John L. JR. Schroder

INTRODUCTION-SYSTEM DESIGN The design of an underground mining operation requires the integration of transportation, ventilation, ground control, and mining methods to form a system which provides the highest possible degree of safety for mine personnel; the lowest cost per ton of product; the most suitable quality of final product, considering its intended use; the maximum percent recovery of the reserve, consistent with cost and subsidence rights ; and the minimum disturbance to the environment. The selection of mining method is generally the controlling factor in overall mine design. Some important considerations in the choice of mining method are : 1) Depth of cover over the coal seam. 2) Strength and thickness of rock strata in the overburden. 3) Characteristics of the coal seam, including thickness, nature and location of impurities or irregularities, dip (or pitch), nature and direction of cleavage, hardness, and structural strength of the coal. 4) Nature and strength of the immediate roof (top) over the coal seam. 5) Nature and strength of the floor (bottom) under the coal seam. 6) Presence and amount of explosive gas (methane). 7) Quantity of water likely to be encountered in the mining operation. 8) Any previous or concurrent mining in seams above or below the coal seam to be worked. 9) Any restrictions on recovery required for subsidence control. Topographical and geological reports and maps of the area are usually available from state and federal geological surveys, and contain excellent general information to guide more detailed investigations. Exploratory drilling, and the plotting and analysis of the information thus obtained, are essential steps in any development program, and aid in the correlation and evaluation of information gained by other means. Other good sources of information are operating coal mines that may be in the same

Jan 1, 1981

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Jan 1, 1940