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By K. -A. Björkstedt

INTRODUCTION Strassa lies in the central part of Bergslagen, a tradi¬tional mining district, on the eastern side of the Stora Valley at an elevation of about 200 m above sea level. A railway siding runs between the mine and the Stora railway station from which there are railway connections to the shipping port and iron and steel works in Oxelosund, about 224 km away. The distance to the provin¬cial capital Orebro is about 60 km. The climate is typi¬cal for this part of central Sweden and is illustrated by the diagram of monthly precipitation and temperatures for the years 1968-1975 (Fig. 1). HISTORY There is no certain information as to when the Strassa mine was first worked, but it is known from sur¬viving accounts of mine inspectors that there were smelt¬ing works in operation in nearby villages in the 12th century. An example is the Gusselhytta ore smelting works, 10 km south of Strassa, which dates from this period. Around the year 1540 there were two smelting works in Strassa, the Upper Karberg and Lower Karberg works. Ore for these smelters was probably taken from Strassa and from the adjacent Blanka mine. In the year 1624 Strassa is mentioned by the painter Jons Nils Krook in an account of the iron mines in the Linde mining district (Linde Bergslags Jarngruvor). Several mines were listed in the area, the deepest being about 30 m. An impressive power installation is mentioned in 1639, including a piston system of lashed poles for transmit¬ting power from the Stora River to the Strassa fields. Its length was 2670 m. Common ground comprising about 20.2 km2 (5000 acres) of forest was allocated in 1689 for the furtherance of mining operations. Until the beginning of this century only the rich cen¬tral parts of the ore body were mined and these yielded, after handpicking, lump ore suitable for smelter feed. An example of the ore grades from these early times is an analysis of ore from the "Big Mine" (Storgruvan) from the year 1873: 48.5% Fe, 0.008% P, and 0.06% S. This same year a total of about 18 000 t was ex¬tracted from the Strassa mine. OWNERSHIP The mine was owned and run until 1874 by a min¬ing association made up of 119 so-called "bergsman," who were homesteaders often engaged in agriculture and timber-cutting as well. In that year the Strossa Grufvebolag (Mining Co.) was founded. In 1906 it was con¬verted into a joint stock company, the Strossa Gruveaktiebolag. This was acquired in 1907 by Metallurgiska AB for the implementation of Gustav Grondal's beneficiating and briquetting methods, for which the Strassa ore was well suited. The same year saw the completion of a new ore dressing plant with an annual production of 46 000 t of ore concentrate. In 1911 the mine passed to new hands, and in 1913 it was purchased by an Austrian company. Extensive new installations were made and in 1915 a new dressing and briquetting plant was completed with twice the capacity of the old one. In 1917 the Strassa mine was acquired by Granges. Be¬cause of unfavorable business trends and technical diffi¬culties, mining operations were brought to a close in 1923. Pumping kept the mine free of water until 1933 but it was completely filled ten years later. Up to 1950 the surface buildings and installations remained intact but the large dressing and briquetting plant burned to the ground in that year. Today only the machine shop re¬mains from this earlier period of operation, now housing parts of the Mineral Processing Laboratory. The decision to take up mining operations again was made in 1955 and construction work began the follow¬ing year. Of the old installation, only the "southern shaft" could be used for some development drifting after it had been completed with a new headframe. Other¬wise, all the buildings and installations required for the operations had to be rebuilt. New installations ready by 1960 were office and personnel facilities, a new shaft and headframe, a sorting and concentrating plant, a macadam plant, settling basins, pump stations, and a railway and yard with transport equipment. The instal¬lation was completed with two plants

Jan 1, 1982

By L. A. Rapoport, W. J. Leas, C. W. Carpenter

A program of scaled flow model experiments has been undertaken to study the performance of five-spot water floods. The modeling procedures are discussed and the construction and operation of the flow models are described. The tests were specifically designed to investigate the effects of injection rate and of oil-to-water viscosity ratio on oil recovery. The areal sweep of five-spot water floods was also studied, and direct comparisons between five-spot and linear flooding behavior have been established to permir better utilizatfon of core analysis results for purposes of field waterflood evaluations. INTRODUCTION For a long time reservoir engineering procedures have been confined to the consideration of idealized, one-dimensional or linear systems.11* However, direct application of the linear waterflood mechanism to the evaluation of actual field performance is justifiable only for the purposes of rough approximation. To obtain a realistic analysis of field behavior, it is necessary to account for reservoir geometry and well distribution, and to take into consideration the configuration of the flow of fluids between injection and producing wells. In the past it has been attempted to account for the complex nature of the fluid movement in the reservoir by associating each well pattern with a certain "sweep-efficiency constant", e.g., a sweep efficiency of 72 per cent for the conventional five-spot pattern.' More recently, methods of waterflood evaluation have been improved by the introduction of variable sweep-efficiency factors, defining the areal coverage as a function of cumulative water injection and viscosity or "mobility" ' ratio.""' These variable sweep-efficiency factors, derived from miscible fluid displacement experiments, require the assumption of piston-like displacement of oil by water. This implies that only one fluid, either oil or water, is moving at any one point of the reservoir ht any moment. Such an assumption, however, is incompatible with the bulk of the experimental evidence and also theoretical considerations which indicate that usually in the area invaded by water, both oil and water are flowing in varying proportions, in accordance with the relative permeability characteristics of the porous medium. Another recent approach toward the evaluation of five-spot flooding behavior involves correlations of the areal sweep efficiency (observed in immiscible fluid experiments) with a "mobility ratio," based on the consideration of the "average" relative permeability to the displacing fluid. Combination of these correlations with a modification of the Buckley-Leverett frontal-drive equation leads to a computational procedure for predicting the recovery performance of areal water floods. While of considerable interest, this approach has no complete theoretical justification and, therefore, does not entirely clarify the features of five-spot water floods. Rigorous, theoretical treatment of waterflood behavior in two- or three-dimensional systems as encountered in the reservoir, including proper consideration of relative permeability and capillary-pressure functions, is as yet beyond the scope of analytical or computational methods. The, behavior of such water floods can, however, be studied by means of scaled flow model ekperiments, which is the approach followed in these investigations. The flow model experiments described in this paper simulate a uniform, horizontal formation of constant thickness with fully penetrating wells. The system under consideration is entirely liquid saturated and not influenced by compressibility effects, which is representative of the situation achieved after gas fill-up under practical conditions. Gravity segregation effects are assumed to be negligible, which implies a relatively thin reservoir formation and/or low vertical permeability. THEORETICAL CONSIDERATIONS For a flow system of uniform thickness with no appreciable effects of gravity segregation, the analysis of water-oil displacement can be confined to a two-dimensional treatment based on the consideration of a "slice" of unit thickness. Referring to previously established derivations," the displacement of oil by water in such a system can be described by the following, dimen-sionless equations:

By E. L. Thyse

Keywords: Pyrometallurgy, converter, nickel, matte, distribution Metallic elements occupy a wide range in granulated nickel converter matte phases. The distribution of these metallic elements, particularly during crystallization path events, appears to be poorly understood. The present study serves to investigate the distribution of Fe, Co, Ni, Cu, and Pb between nickel-sulphide, copper-sulphide, and alloy phases. A novel combination of thermo-chemical modelling, electron microscopy, and electron probe microanalysis was applied to understand crystallization path events and the associated metallic element distribution. This study has demonstrated that major Ni incorporated during copper-sulphide crystallization would serve to mobilize minor Co towards distribution in copper-sulphide phases. In addition, solid solution substitution of Cu by Ni is manifested by compositional zoning of copper-sulphide phases. The role played by Ni during alloy crystallization suggests that minor Co was also mobilized, along with an increase in principal Ni. Fe, by contrast, seems to be incompatible with principal Ni. In addition, solid solution substitution of Ni with PGEs and Cu is manifested by compositional zoning of NiCu-alloy phases.

Jan 1, 2011

By P. E. Richardson, N. P. Finkelstein, Z. Chen

The large band gap and insulating properties of sphalerite have limited the application of standard electrochemical techniques for studying its reactions with flotation reagents. This problem has been overcome by developing electrodes that rely on the surface conductance produced when sphalerite is activated with copper or other heavy metals. A brief exposure of the mineral in copper sulfate solutions results in electrodes that exhibit significant electrochemical activity characteristic of copper sulfides. Voltammetry studies conducted on sphalerite activated in 10-4 M CuS04, solutions at pH 4.6 provide information on both the amount and the nature of the activation product(s). It is shown that copper activation is strongly dependent on the electrochemical potential applied during activation. Activation at potentials increasingly negative of the rest potential causes a progressive increase in the amount of copper on the surface; activation at potentials positive of the rest potential decreases the amount of copper on the surface. The results indicate that both the kinetics of copper uptake and, possibly, the chemical composition (stoichiometry) of the activation product are controlled by the electrochemical potential.

Jan 1, 1995

By Oswald J. Heinrich

The responsibility resting upon the owners and managers of mines where fire-damp is generated, renders it a matter of imperative duty that a frill and correct statement of any explosion that occurs should be given to the public. But such a statement must be based on the testimony of impartial eye-witnesses or trained experts. Then a sad experience may result in the adoption of effective preventive measures. In the absence of any official and professional body of mining experts in the State of Virginia, the writer desires to submit to his professional brethren of the American Institute of Mining Engineers an account of an explosion which recently took place in a colliery under his management, and earnestly desires a thorough and impartial discussion on the Fats submitted. Description and Plan of the Mine.—The "Grove" shaft is situated at the nearest point 825 feet from the last old workings upon the Midlothian property, south of the old Pump shaft works. (See Plate 111.) The shaft was sunk by the former company, some thirty years ago, to the depth of 622 feet. An incline over 230 feet long, starting in a northwest direction at a depth of 105 feet above the shaft bottom, had been driven to explore the coal on the dip. A return of very small dimensions had been driven round the shaft on the south side, entering the upcast chamber. With the information obtained from the former company that a seam of coal about four feet thick, at a depth of 490 feet, had been found, and with the information obtained by the present proprietor, R. T. Burrows, of Albion, N. Y., by means of two boreholes on the dip west of the old shaft, in both of which workable seams of coal at various depths had been ascertained, the clearing out and thorough retimbering of this shaft were commenced in March, 1873. Borehole No. 1, at a depth of 4773 feet from surface, revealed a seam of coal. This borehole was afterwards carried to a depth of 1140 feet to the granite, without finding any other seam of coal. Borehole No. 3, at a depth of 608 feet, revealed the first coal-seam, 14 1/2 feet thick; at 633 feet, the second, 12 feet thick; at 662 feet the third, 1 foot thick, and at 692 feet the fourth, 43 to 5 feet thick. The borehole was stopped at a depth of 715 1/2 feet.

By Oswald J. Heinrich

THE responsibility resting upon the owners and managers of mines where fire-damp is generated, renders it a matter of imperative duty that a full and correct statement of any explosion that occurs should be given to the public. But such a statement must be based on the testimony of impartial eye-witnesses or trained experts. Then a sad experience may result in the adoption of effective preventive measures. In the absence of any official and professional body of mining experts in the State of Virginia, the writer desires to submit to his professional brethren of the American Institute of Mining Engineers an account of an explosion which recently took place in a colliery under his management, and earnestly desires a thorough and impartial discussion on the facts submitted. Description and Plan of the Mine.-The "Grove" shaft is situated at the nearest point 825 feet from the last old workings upon the Midlothian property, south of the old Pump shaft works. (See Plate III.) The shaft was sunk by the former company, some thirty years ago, to the depth of 622 feet. An incline over 230 feet long, starting in a northwest direction at a depth of 105 feet above the shaft bottom, had been driven to explore the coal on the dip. A return of very small dimensions had been driven round the shaft on the south side, entering the upcast chamber. With the information obtained from the former company that a seam of coal about four feet thick, at a depth of 490 feet, had been found, and with the information obtained by the present proprietor, R. T. Burrows, of Albion, N. Y., by means of two boreholes on the dip west of the old shaft, in both of which workable seams of coal at various depths had been ascertained, the clearing out and thorough retimbering of this shaft were commenced in March, 1873. Borehole No. 1, at a depth of 477 1/2 feet from surface, revealed a seam of coal. This borehole was afterwards carried to a depth of 1140 feet to the granite, without finding any other seam of coal. Borehole No. 3, at a depth of 608 feet, revealed the first coal-seam, 142 feet thick; at 633 feet, the second, 12 feet thick; at 662 feet the third, 1 foot thick, and at 692 feet the fourth, 4 1/2 to 5 feet thick. The borehole was stopped at a depth of 715 1/2 feet.

Jan 1, 1877

The Kunwarara magnesite deposit is the largest known magnesite (MgCO3) deposit in Australia. It is located under black soil plains to the north of Rockhampton in Queensland. The deposit consists of magnesite nodules set in a fine-grained sandstone, siltstone and mudstone matrix. The magnesite is derived from weathering of serpentinite in adjacent hills and transport of magnesium in solution to a Tertiary/Quaternary river system. The magnesium has come out of solution in the active river system, and combined with atmospheric and biogenic CO2 to form magnesite nodules. The magnesite can be very pure with certain varieties up to 98 per cent MgO (LOI free). The deposit has been mined since 1991 and processed into magnesia (MgO) which is sold into refractory, agricultural and other markets. The geologist is an integral part of the small mine technical team. The role of the geologist includes characterisation of the ore through drilling and bulk sampling, supervision of geological aspects of mining and assistance in the optimisation of ore processing.

Jan 1, 2000

By John D. Roach

DURING an investigation of the yttrium-hydrogen system aimed at producing solid yttrium hydride specimens containing various amounts of hydrogen, it was observed that yttrium containing approximately 2 wt pct H exhibited a tendency to crack and crumble to a powder on standing in air at room temperature. It was also observed that longer hydriding times, at a given temperature, increased this susceptibility to powdering without an increase in hydrogen content, also that moisture in the air is necessary for the observed effects to occur. Even if cracking did not occur, there was a continual formation of a light gray powder on the as-treated surface of the hydrided yttrium and this reaction continues until the entire hydrided piece has been reduced to a powder. Storing the hydrided specimens in the absence of air or removing the surface either by machining or grinding were effective means of preventing this disintegration of hydrided yttrium. In an attempt to determine the reason for this powdering phenomenon, the surfaces of a number of as-hydrided yttrium specimens were examined by X-ray techniques. In all cases the X-ray pattern obtained showed the major phase to be yttrium hydride (YH2) as would be expected since the specimens contained 2 to 2.2 wt pct H. In some cases a trace of yttrium oxide was observed. There was also a third phase present on the surface of these hydrided specimens which could not immediately be identified. This unknown phase was a face-centered cubic material, NaC1-type structure, with a lattice parameter of 4.855 and a calculated density of 5.914 g per cc. Very slight hand polishing of the surface of the hydrided yttrium specimens completely removed both the unknown phase and the traces of oxide so that only the yttrium hydride pattern remained. Based on the X-ray patterns the quantity of the unknown phase on the surface of the hydrided yttrium appeared to be directly related to the susceptibility of the material to powdering. Work at the Denver Research Institute on the yttrium-nitrogen binary system showed that yttrium nitride (YN) is a face-centered cubic material with a lattice parameter of 4.878A and a density of 5.890 g per cc. They also noted that this compound rapidly disintegrated to a powder on standing in air. The unknown pattern observed in the above specimens corresponds very closely to that of the nitride—fcc structure, 4.885 parameter, and density of 5.914 g per cc. The presence of this thin film of nitride on the surface of the hydrided specimens probably accounts for the observed powder formation and crumbling. The nitride reacts with the water vapor in the air (verified by private communication from Dr. C. Huffine, General Electric Co.) to yield yttrium oxide and ammonia. Ammonia is readily detected when hydrided yttrium specimens were allowed to stand in bottled moist air. The powder formed on the surface of the hydrided specimens was shown by X-ray analysis to be yttrium oxide. This reaction appears to occur primarily at the grain boundaries since discrete particles of yttrium hydride separate from the specimens during this powdering process. The reaction of yttrium nitride with water vapor is believed to be as follows: 2YN + 3H2O - Y203 + 2NH3. Despite the fact that this nitride is present only as an extremely thin surface film, if the above reaction is not prevented by removing this film from the surface either by machining or grinding, the reaction continues until the entire hydrided piece has been reduced to a powder. To account for this continuation of the reaction, it is believed that the following reactionalsooccurs: 2YH2 + 2NH3 -2YN + 5H2. The nitride produced by the latter reaction reacts in turn with water vapor. The reaction therefore becomes autocatalytic and continues until the hydride has been consumed and the entire piece reduced to oxide powder. The amount of nitride required to initiate this reaction is quite small and this nitrogen contamination can occur from a number of sources—hydrogen gas employed, minute leak in the hydriding apparatus or even from degassing of the reaction vessel itself. Longer hydriding times increase the possibility of nitrogen contamination and this is especially true when a dynamic gas system is employed in the hydriding process. The production of stable, solid hydrided yttrium is dependent on the complete absence of nitrogen contamination during processing. If such contamination does occur, powdering of the hydrided product can be prevented by removing the nitride from the surface e.g. grinding or by preventing access of air to the pieces e.g. sealing in wax or plastic. The author wishes to express his appreciation to General Electric Co. for sponsoring this research and for permission to publish the results of work under Subcontract AT-93.

Jan 1, 1962

By J. S. Marsh

OPPORTUNITY to pay tribute to the memory of Professor Henry Marion Howe is a strenuous assignment as well as an honor. Upon recalling Howe lecturers and lectures of the past 25 years, glancing over the list of those earlier, and rereading Howe's books, I arrive at several conclusions: 1—Many lecturers either worked under or knew Professor Howe. 2—It is virtually impossible to pick a subject on which Professor Howe did not touch. 3—There is precedent for a technical paper based upon pursuit of a single subject. 4—There have been listening lectures and reading lectures. There is solid comfort only in 2: the subject field is wide open. I did not know, nor even ever saw, Professor Howe, so can supply no fitting reminiscence. As a college student I was dimly aware that he counted among the giants. Fuller appreciation of his stature came with reading his books and papers, growing acquaintance with some of his associates, and the intrinsic dignity of the climax of the Annual Meeting, beginning at four o'clock of a Thursday afternoon in the auditorium of the Engineering Societies Building in New York. As for producing the technical paper sort of thing, it is my lot to have reached an age and assignment such that to do so would be to filch information from those who did the work and whose story is theirs to tell; for this I have no enthusiasm. As for the final conclusion, Professor Howe was one of the chosen few so highly expert at expository writing that he could produce a lecture or paper that reads as though it would also have listened well. One of his tricks was the free use of words not ordinarily part of the technical vocabulary, provided that such words were likely to communicate most precisely what he had in mind. How wonderful it would be for all who must read reports by the ton if ability at exposition could be taught with the effectiveness open, say to, differential calculus! Perhaps Professor Howe should be required college reading even if for no other reason than to prove that technical writing need be neither dull nor diffuse. My assignment is clearly still strenuous. Another point to consider is the fact that metallurgy is now so tremendously diversified that hope of finding a topic of universal appeal is negligible, even if one were competent enough to be permitted free choice. That which follows is, therefore, a compromise composed of necessity and of the obligation to attempt to avoid boring to slumber those of you who are not especially interested in the general subject chosen. The Iron and Steel Div. is now essentially a process metallurgy division, heavily concerned with the smelting of iron and the making of steel. The American Iron and Steel Inst. figure for present steel capacity of this country is 125,828,310 net tons; how this is divided among processes is indicated by the production totals for 1953, shown in Table I. The glamor girls and boys make the front page and so it is with steelmaking processes. If there is an Antarctic Daily Bugle, it undoubtedly has carried stories of revolutionary development, such as oxygen processes and vacuum melting, and stories of the incomparably rosy destiny of electric arc melting. All such certainly have their place and their future; meanwhile, it is the sturdy and old reliable open hearth that accounts for the bulk of production reported back on the financial page, and it is the old reliable that is most likely to continue to account for the bulk for some perfectly sound raw material, technologic, and economic reasons. This, plus the fact that next year marks a centennial (for it was in 1856 that Frederick and William Siemens conceived the regenerative open hearth), is reason enough to talk about open hearth furnaces, but is not the real one. The real reason is that in some years of association with open hearths, I have accumulated—in addition to a genuine liking and respect for them—certain odds and ends of fact and fancy that this lecture provides a unique chance

Jan 1, 1956

By H. H. Rachford, D. S. Howard

In the withdrawal of fluids from tilted aquifers it is of value to be able to predict pressure patterns during the course of the pressure decline. As an example of this, in the displacement of fluids from oil reservoirs being produced by water-drive, prediction of the reservoir pressure is an important factor in fixing production rate. The prediction of pressure patterns in the depletion of an inclined aquifer has conventionally been made by solving the well known flow equations for the pressure Patterns in a horizontal aquifer, and correcting these patterm for the pressure difference due to the hydrostatic head associated with the vertical distance between a point in the horizontal aquifer and the corresponding point in the tilted aquifer. Inasmuch as the equation describing flow in tilted strata is different from that for horizontal flow. there was question as to the validity of the simple corrective procedure current1y in use. To examine the magnitude of error made using this procedure, an inclined aquifer was chosen for which the error was expected to he large, and the equation for flow in this tilted stratum was solved numerically. The results of the conventional procedure of correcting the solution of the equation for horizontal flow was compared with the solution of the correct equation by studying the difference between predicted pressures by the two methods. Three forms of the hydrostatic head correction were studied. The results for all three forms of the head correction are similar. The error in the approximate solution is quite dependent on the position of the horizontal aquifer with respect to the inclined one. The error is greatest at the producing end ad increases with amount of fluid withdrawn. The magnitude of the errors is not large if a good choice is made for the position of the horizontal datum. The most satisfactory position for the datum in all cases is near or slightly about the mid-point of the inclined stratum. Use of the best approximate method studied with the best datum tested yielded predicted pressures in error no greater than 7 psi. INTRODUCTION The movement of water by expansion from an aquifer to displace fluids from a contiguous oil reservoir was first recognized over 20 years ago.9 Since that time much work has been done to relate this mechanism to reservoir behavior mathematically. The result of this effort on the part of many workers has led to an understanding of underground fluid movement which has enabled us to follow and predict the behavior of the water-drive reservoir. One reason that good progress has been made in the application of mathematics to the study of pressure patterns and water movements in aquifers has been that the differential equation which describes this process is of a relatively simple form. The equation for the isothermal, horizontal, viscous flow of water through homogeneous sandstone is identical with that for the diffusion of heat through homogeneous solids, and this equation and means of obtaining solutions for it with many types of boundary conditions have been studied extensively. That this work has been highly successful is clearly emphasized by the large number of solutions developed for problems of interest both by methods of formal mathemat-

Jan 1, 1957

By R. I. Jaffee, F. C. Holden, H. R. Ogden

Ti-Mn alloys were studied in order to determine the factors affecting the mechanical properties of &stabilized titanium alloys. The principal compositional factors have been found to be solid-solution strengthening, the martensitic transformation, and instability of the P phase. Structural factors, such as grain size and shape, were found to have more influence on ductility and toughness than on strength. THE alloys of titanium with the &stabilizing elements offer the chief hope for developing useful heat treatments. Conversely, the heat-treatment response possible with the p-stabilizing elements causes difficulties such as weld embrittlement and thermal instability during service at elevated temperature. Therefore, it is of considerable technical importance to understand the factors that govern the heat-treatment responses in these alloys, so as to be able to soften the alloys if they are embrittled by an adventitious heat treatment, to render them stable in service, or to harden them while maintaining adequate ductility and toughness. The Ti-Mn system is a good example with which to demonstrate the factors that govern strength and heat-treatment response in. a P-stabilized system. The region of the a-j3 transformation, after the work of Maykuth, Ogden, and Jaffee,' is shown in Fig. 1. The a solubilities are low. This means that, in the two-phase a-p field, partition of manganese between the two phases occurs practically only to the P phase. The solid-solution effects are therefore chiefly the result of solution in the ,8 phase. The eutectoid occurs at 20 pct Mn and 550°C, and proceeds very sluggishly. It can be ignored in heat treatments in hypo-eutectoid alloys, except for long-time storage at low temperatures below the eutectoid temperature. Even here, there is some question that diffusion proceeds far enough to involve eutectoid decomposition. Three compositional factors bear on the mechanical properties and structure of titanium a-P alloys. These are: 1—solid-solution strengthening, 2—mar-tensite transformation, and 3—P instability. Combined with the purely structural factors of grain size and shape, these govern the properties of the alloys. Perhaps the most important factor is solid-solution strengthening. The facts that manganese dissolves in a titanium to a maximum of only 0.5 pct and most of the manganese partitions to the p phase dominate the structure and properties of the a-/3 alloys. Because of the overwhelming preference of manganese for the ,3 phase, the P phase is harder and the a phase is softer. The relative amounts of both phases may be modified by heat treatment in the a-P field as application of the lever rule in the two-phase field clearly shows. A critical temperature in the Ti-Mn diagram is 800°C. This is the P transus temperature of the 6.4 pct Mn alloy, which has the lowest manganese content which will form all retained-P phase on quenching to room temperature. Any quench performed from below 800°C will produce a structure consisting of only a, a plus retained P, or retained p, for alloys of the compositions used in this work. For alloys containing less than 6.4 pct Mn, any quench from above 800°C will always result in structures

Jan 1, 1955

By N. Shariatmadari

In recent years, determining bearing capacity of piles from in-situ testing data as a complement of static and dynamic analysis has been used by geotechnical engineers. In this paper, different approaches for estimating base bearing capacity of piles from SPT data has been compared and evaluated. A new method for pile base bearing capacity in sand is proposed using SPT records. Data averaging, failure zone extension, and plunging failure of piles have been addressed in the proposed approach. A data bank comprising 10 full scale static load tests on instrumented piles and 26 dynamic tests, analyzed with signal matching technique by CAPWAP software, was used for validation. The SPT data that was performed close to pile locations is also included in the data base. Comparison of current methods by error investigation with Log-Normal approach demonstrates that the proposed method predicts pile base capacity with more accuracy and less scatter in comparison to other methods.

Jan 1, 2006

By William J. Ferguson

At Henderson Operations the 24-kilometer conveyor system consists of three main conveyors with lengths of 1.2 km, 16.7 km and 6 km. Upon commissioning of the conveyor system, the Henderson maintenance department was challenged with the inspection and monitoring of over 50,000 idlers. To take a proactive approach to idler maintenance, an idler tracking database system was developed. Using the idler tracking program, the Henderson Crush & Convey department now has the ability to provide data for engineering solutions to idler problems and to manage thousands of idlers for priority replacement work and warranty purposes. Recognizing the potential for injury to Henderson employees in the process of idler handling, Henderson approached idler replacement as a material handling process. Material handling solutions have resulted in failed idlers being replaced and discarded from the underground environment without ever touching a human hand. This paper will detail how Henderson tracks idlers in all primary conveyor belts and how an innovative idler handling system can result in reduced risks to conveyor maintenance employees.

Jan 1, 2004

By R. W. Raymond

PROF. HENRY Louis, of Armstrong College, Newcastle-on-Tyne, England, a distinguished member of this Institute and other technical societies, has recently sent to the Institution of Mining Engineers, and to the Institution of Mining and Metallurgy, in Great Britain, a communication, the substance of which is as follows Prof. Louis calls attention to the absence of any recognized English term for the definite description of a frequent phenomenon-namely, the oblique position of an ore-body in the plane in which it. lies. In this connection, Prof. Louis says that in America the term " pitch " has been occasionally applied to this relation, and proposes that, in the literature of ore-deposits, it shall be hereafter restricted to this particular meaning, the angle of the pitch, like that of the dip, being always measured from the horizontal, and the dip being always taken at right angles to the strike, while the pitch is taken on the plane of the dip, in the direction of the strike. Thus, for example, in a vein or bed striking N-S., and dipping, say, 45° E., there might be an ore-body, the axis of which pitched northward, at an angle of 45° below the horizontal line, as measured on the plane of the inclosing vein or bed. If the term " pitch " had this universally recognized meaning, the situation could be simply and perfectly expressed by its use, in connection with the general data of strike and clip. As Prof. Louis says ?This suggestion has a certain practical value, and is not merely academic. Every mining engineer of practical experience in such types of ore-deposit will have met with cases, in which a vertical shaft has been sunk with the object of cutting the deposit, but has missed it, because the dip alone has been taken into account, while the pitch has been overlooked." He adds the acute observation that such an oversight is much likelier to occur with regard to a condition not definitely desig-

Mar 1, 1908

By M. Nusheh, C. Bumby, A-H. Puichaud

"Vanadium-bearing titanomagnetite ironsand deposits occur along a 480 km stretch of the west coast of New Zealand. In the present study, the leaching behaviour of NZ ironsand was investigated in hydrochloric acid (2–10 M) solution at 50–80°C for 2 h. The resulting leaching liquors exhibited Fe/V ratios in the range 150–180. The distribution of trace element contaminants between solution and solid residue was investigated, and processing conditions have been determined for rapid and efficient dissolution of Fe, Ti and V into the leaching solution. In addition, mineralogical analysis of raw ironsand and leached residues was performed in order to study the leaching mechanism. Samples were characterized by scanning electron microscopy, energy dispersive X-ray spectroscopy (EDX), X-ray fluorescence spectroscopy, qualitative and quantitative X-ray diffraction and inductively coupled plasma mass spectrometry (ICP). Results indicate that the dominant dissolution process follows shrinking particle mechanism. However, ironsand particles containing ex-solution lamellae exhibited a different leaching behaviour compared to homogenous titanomagnetite particles.INTRODUCTION Titanomagnetite ironsand (blacksand) forms Quaternary onshore beach and dune deposits and offshore marine deposits (Figure 1) along 480 km of coastline from Kaipara Harbour south to Wanganui on the west coast of the North Island (Christie, 2016). It attracted the attention of scientists from the earliest times (Galbreath, 1998). The principal minerals in ironsand are titanomagnetite (TTM), ilmenite, quartz and ferromagnesian minerals such as augite, hornblende, diopside, and olivine with smaller amounts of rutile and apatite (Mauk, 2016).Several large ironsand deposits have been discovered and exploited since the late 1960’s. The Waikato North Head and Taharoa mines are currently operating and they have a 2015 measured resource of 400 million tonnes in total and provide over 2 million tonnes of concentrate per year by simple cost effective magnetic separation process (Christie, 2016; Christie & Brathwaite, 1997). The concentrate contains not only 7–8% TiO2 and 56–60% SFe, but also 0.3–0.6% V2O5, is both a valuable titanium mineral and a source of vanadium. All the production is used for steelmaking industries. The other potential for utilizing more of this resource without requirement of integration into an iron and steel producing plant is to develop a hydrometallurgical route with advantages of higher efficiency along with lower emissions. Currently, there is no existing commercial process capable of extracting all three of these elements as industrial-grade commodity products. Acid leaching of TTM concentrate with higher degrees of recovery and selectivity is a key step in any possible hydrometallurgical route."

Jan 1, 2017

IT will be noticed, from the following contributed papers on the various plant, what an infinite number of parts are required to be kept in active operation in order to derive a resultant product of commercial value from the ores and materials received at the works. Crude ores and concentrates, in every degree of complex structure and resistance to treatment, are arriving daily from near and distant places, moreover their metallic contents are so varied that an outsider, unacquainted with the operations, would naturally wonder how it is possible that the rare elements are to be sorted out and classified into their true order, to meet industrial requirements.The proposition is necessarily not a simple one and owners of mining properties have naturally refrained from smelting their complex ores on account of the technical difficulties of attaining high efficiency and recovery of the metals, where limitations are imposed in procuring a variation of ores that lead to the best smelting and refinmg result. This difficulty obviously imposes many disabilities on the smelter, who is prepared to accept all such complex ores, and it may be said that were therenot available a large weekly supply of lead concentrates from the company's own mine at Broken Hill, of standard quality, much of the material being treated in the world today would certainly not be marketable.In order to sustain full efficiency, each section of the works must maintain its.true balance and synchronize with the requirements of such other departments as are dependent on its partially treated materials. This harmony of operations is not only imperative as regards tonnage output, but also relies on the condition of the resultant treatment by each plant, no large variations being permissible.The various officers are, therefore, dependent on one another to keep the order of the operations flowing regularly. This is accomplished by constant intercommunication and observation, as...

Jan 1, 1918

By A. Alonso, J. Daubenspeck

In the nickel extraction process at Nicaro, Cuba, dried and ground nickel bearing laterites are reduced at 1300-1400F in multiple-hearth furnaces. Nickel is leached from the reduced ore by aeration in an aqueous solution of ammonium carbonate. Nickel carbonate is precipitated from the liquor and calcined to yield nickel oxide. The following are the more important modifications of the Nicaro process since plant rehabilitation and expansion. Ten percent increase in tonnage was achieved without decreasing the fineness of the ore by minor mechanical modifications of the grinding plant. Another five percent increase in tonnage was attained by intentionally overloading the ore prepration and leaching circuits in an effort to control cobalt extraction. Modifications at the sintering plant allowed it to handle the entire output of the expanded plant. A fifty percent increase in leach liquor cooling facilities alone provided an increase in nickel production of five percent. Stronger fresh leach liquor for increased nickel washing efficiency was obtained through improved distribution of water to the ammonia absorption towers. Advantage was taken of the reduced calcine's capability to precipitate cobalt, and modifications in the leach liquor circulation provided controlled cobalt solubility with minimum harm to nickel extraction. Several process changes at the drying plant brought about a ten percent reduction in the fuel oil consumption at this pint. Controlled air addition to the reduction furnaces' upper hearths burned formerly wasted combustibles thereby reducing oil consumption by twenty percent. The nickel extraction process successfully emplayed on the extensive but lowgrade lateritic ores of Eastern Cuba at the U. S. Government owned Nicaro plant, has been previously described in various publications. The large mine and plant operated by the Nickel Processing Corporation, located on the northeastern shore of Oriente Province, have been recently expanded to a yearly capacity of about 50 million pounds of nickel. Essentially, the operation involves open pit mining of a limonite-serpentine ore mixture containing 1.3 to 1.4 percent nickel and 34 to 38 percent iron. The comparatively lumpy, wet (25 to 30 percent moisture) ore is first dried and ground prior to chemical treatment, to which the entire 7,000 tons per day feed must be subjected. In the initial chemical action the dried and ground ore is reduced with carbon manoxide and hydrogen at 1300-1400° in multiple-hearth furnaces. The resulting metallic nickel is selectively leached from the reduced calcine by aeration in an

Jan 1, 1961

By Robert J. De Angelis, James H. Barker

An improved method for solving dynawzical dislocation problems using a digital computer is described in this paper. Interactions between two distinct types of dislocations were studied: attractive screw dislocations; and Lomer lock forming dislocations. One dislocation is positioned in the lattice and is initially at rest, while the other dislocation is moved through the lattice on an intersecting slip plane at a constant velocity in the range 0.5 to 0.999C. (C is the transverse velocity of sound.) The results obtained from these computations indicate that screw dislocations account for a small fraction of the total strain over a wide portion of the range of velocities studied. They further indicate that mixed dislocations mainly repel other dislocations in the neighborhood of the active glide plane. From this a possible explanation for cell formation is put forth. The density of Lomer locks expected to exist after a strain of 0.2 was found to be 1.4 x 106 cm-2 which is in good agreement with indirect experimental estimates. IN the past, predictions of favorable or nonfavorable dislocation reactions were based on the associated changes in elastic strain energy. Such considerations take no account of the probability of the two dislocations coming into contact to react. Venables1 was the first to approach these probabilities by considering the interactions between two moving screw dislocations on perpendicular glide planes. Because of the restrictive types of dislocations and glide plane geometry employed, his results have limited application to metallic crystals. The work to be presented here develops a general approach to solving dynamical dislocation problems; either dislocation-dislocation interactions, presented here in detail, or dislocation interactions with any other suitably defined stress field. Two types of dislocation-dislocation interactions common to face centered cubic (fee) materials are considered: those between pure screw dislocations of opposite sign on intersecting slip planes and those between mixed dislocations on intersecting slip planes, that can react to form a perfect dislocation. This latter reaction, referred to as the Lomer reaction, produces a locked product dislocation that finds it energitical favorable to disassociate into two Shockley partials and a stair-rod dislocation. This partial configuration known as a Lomer-Cottrell (L-C) lock plays a major role in work hardening of fee crystals. seeger2 names the L-C lock as the prime contributor to Stage II hardening while Kuhlmann-wilsdorf3 and Meakin and Wils- dorf4 also state that it is a significant contributor to work hardening. However, with a few notable exceptions,5-7 direct observations of the Lomer lock and the L-C lock by electron transmission microscopy are scanty, and even these are subject to other interpretations.5,6 In a study of partial dislocations present in austenitic stainless steel, whelan8 did not observe any L-C locks at the head of pile-up groups. This result contradicted existing work hardening theories and led him to postulate an alternate theory based on the stress required to break away dislocations intersecting a pile-up group, from their stacking fault nodes. Due to the importance of the Lomer reaction in producing L-C locks which are an essential feature in current work hardening theories and because there exist no data giving direct quantitative values for the density of locks, and because there has even been some doubt expressed as to whether this important reaction occurs at all, a study of the dynamic behavior of the mixed dislocations which form the Lomer lock was undertaken. Due to their ability to cross-slip with relative ease, screw dislocations play an important role in the deformation of fee crystals. For this reason, the second type of reaction considered here is between screw dislocations of opposite sign. In addition, computations in volving screw dislocation interactions are relatively simple, thus providing a convenient check on the cornputational scheme employed. DEFINITION OF PROBLEM The force exerted on a dislocation due to a generalized stress field is given by the Peach and Koehler9 equation: Here t2 and b2 are respectively the tangent and Burgers vectors of the dislocation, and T1 is the stress dyadic defining the local stress field. The stress field may be externally applied or generated internally by the presence of a lattice defect, such as a second dislocation, as is the case in this work. Frank10 has shown that an equivalent momentum, P, of a screw dislocation can be defined by: Here, EST is the total energy of a screw dislocation and ESo is its rest energy. The left side of Eq. [2] is the time derivative of momentum and the right side is the position derivative of the energy due to the dynamical nature of the dislocation. The total energy of a dislocation is the sum of the potential and kinetic energies. Weertman11 has developed the expressions which were used here; these give the potential and kinetic energies of uniformly moving edge and screw dislocations in an isotropic medium.

Jan 1, 1969

By D. R. Rice

"The Smoky Valley Common Operation is managed and 50 percent owned by Echo Bay Mine' s Round Mountain Gold Corporation. The other partners with 25 percent each are Homestake Mining Company and Case Pomeroy. The operation in central Nevada includes an open pit, 3-stage open-circuit crushing train and heap leaching on asphalt pads. Presently, about 18,000 tons per day of 0.039 troy ounce gold per ton are transferred to the leach pads, and a similar tonnage of leach residue is transferred to a waste dump. One million tons of ore are being leached at any one time.Gold and silver cyanide are adsorbed from pregnant leach solution on activated coconut shell carbon granules. Pressure stripping desorbs values from loaded carbon to electrowinning solution.Three electrowinning cells are operated in parallel. Each contains six cathodes. Each cathode includes about 40 troy ounces (2.7 pounds) of Number 2 medium coarse steel wool. The wetted area of each cathode is about 24 by 30 inches. Solution flow is 2 to 3 gallons per minute per electrowinning cell. Average current flow is 25 amperes per cathode. The voltage is controlled in the range of 3 to 4 volts to maintain desired amperage.Twice a week two cathodes are pulled from each cell and drained. Each cathode is split into four portions. These are put in a steel pan coated with WD-40 and dried overnight.Each dried cathode portion is put into a paper sack. Each group of four sacks corresponding to a cathode is weighed together. The typical weight ratio of loaded cathode wool to new cathodes is 9 or 10 to 1. To each sack is added 35 troy ounces of silica, 45 troy ounces of anhydrous borax, 35 troy ounces of nitre, and 5 troy ounces of fluorspar. The sacks are taped shut."

Jan 1, 1986

By Kefton H. Teague

Olivine was first described and named by Warner in 1790. It is an olive green to greenish gray colored magnesium-iron silicate containing a mixture, in solid solution, of fosterite (Mg2SiO4) and fayalite (Fe2SiO4). Ratios between fosterite and fayalite in olivine are somewhat variable, but in general olivine contains at least 85% fosterite. Specific gravity is variable also, depending on the ratios of fosterite to fayalite. As a rule it is in the 3.2 - 3.4 range. All commercial olivine is obtained from the rock dunite, the most basic member of the peridotite group. Dunites, in addition to being composed primarily of olivine, generally contain assessory pyroxenes and spinels (chromite and magnetite). Dunites', when altered by hydrothermal solutions or metamorphism, often contain alteration minerals such as talc, serpentine, chlorite, vermiculite and amphiboles. In the 1930's and 40's considerable research was conducted on olivine for use as a refractory, plant food raw material, magnesium source for metal, and a source for magnesium chemicals. This research met with only limited success.

Jan 1, 1977