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By Arvid E. Anderson, Frank K. Cameron

The weathering of copper-bearing ores with the formation of a water soluble salt and the recovery of the metal by leaching and evaporation precipitation, are processes long known, which have at various times been utilized with more or less commercial success. It is commonly assumed that the technical principles involved are quite simple and their application to a particular case easy. If these assumptions were true, we should anticipate a long list of uniformly successful leaching operations, which is not the case, and one would be justified only in calling attention to either unique or peculiar features of any particular enterprise; whereas, in studying the problems of the recovery of copper from the old workings of the Ohio Copper Co. in the light of available descriptions of other leaching projects, we have not found the assistance hoped for to guide us to the best realization of our own opportunities. It appears desirable, therefore, to call attention to these guiding principles, together with a description of the facts of accomplishment. Justification lies in the fact that the leaching and precipitation of copper at the Ohio Copper Co. mine in Bingham Canyon, Utah, is successful. This success is measured by an average monthly production of 600,000 Ib. of copper at an average operating cost, to date, of 3.85 cents per pound at the company's workings, or 6.32 cents per pound after paying shipping and smelter charges on the finished copper. A total of 17,076,099 Ib. of copper had been shipped to the smelter from the beginning of operations, August, 1922, to April 1, 1925. Probably these results can be bettered with further study and realization of the factors involved and the devising of methods of meeting them. It is now an assured fact that the enterprise will have a relatively long and continuously profitable life and that its reserves will not, nor cannot, be suddenly and permanently depleted. It is particularly noteworthy that the financing of this enterprise has been entirely from its earnings. The few hundred dollars necessary for the first experiments and production were obtained by the sale of used, and no longer needed, equipment from a mill belonging to the company; and all funds since required throughout the growth and development of the enterprise, from a few troughs with a production of 2000

Jan 1, 1926

By L. R. Mabson, F. M. Russell

INRODUCTION Roan Consolidated Mines Limited (RCM) operates five underground mines - Mufulira, Luanshya, Baluba, Chambeshi and Chibuluma (figure 1) . Associated metallurgical plants provide for the operations of milling, smelting two refineries for copper and a cobalt plant. Primarily a producer of high quality copper (242,000 tonnes during 1980) the Company is also the world's second largest producer of cobalt metal. RCM's remaining ore reserves at 31st March 1980 were 275 million tonnes at 2.84% copper. Monthly ore production is in excess of 1.2 million tonnes. Sublevel open stoping was the choice of mining method for the first underground operations on the Copperbelt for ore with dips steeper than 30 degrees. This was for use at the Roan Antelope mine, subsequently to be called Luanshya mine. The sublevel open stoping system was planned based on the methods used in many of the iron mines of Michigan and New York and in the copper mines of Tennessee. Peterson (Ref. 1) gave the following reasons for employing sublevel stoping, reasons which are still valid today: (1) the working place could always be barred and made safe, and the sublevel exit was always at hand; (2) the working conditions were similar day by day and drilling operations were repetitive; (3) pillar spacing can be varied as conditions demand; and (4) stoping operations could be shut down completely or opened up to full capacity at short notice, without affecting grade or causing wastage of reserves. Such was the success of the method that it was adopted as the basic mining method at all the RCM mines in their initial stages. Although cut and fill, block caving, and sub- level caving methods have been used, sublevel open stoping in its many variants still accounts for nearly 90% of total ore production in the Company. The following descriptions outline some of the applications of sublevel open stoping within RCM. LUANSHYA Geology There are two main ore deposits at Luanshya: the Roan (Luanshya Mining Area), in which mining began in 1930, and Baluba, in which mining production started in 1973 (figure 2). Both orebodies occur in synclinal and variably folded basins. These basins are underlain and completely surrounded by much older schists and granites. The ore formation, an argillite, lies near the base of the sediments and is underlain by conglomerates, quartzites and arkoses. Copper mineralisation is mainly found in two well defined zones, separated by a varying thickness of pyrite mineralisation. In the eastern part of the Roan deposit only the upper orebody is present, but west of 18 shaft and at Baluba the lower orebody predominates. Locally the subjacent quartzites of the footwall formation are also mineralised with copper. The copper mineralisation is found princi- pally as disseminated sulphides, mainly chalcocite, bornite and chalcopyrite. Near the surface the sulphides have been changed to oxides. At Baluba the cobalt mineral carrollite is found with the sulphides. Mining methods Sublevel open stoping is the basic mining method. In its standard form this method is best suited to steeply inclined orebodies, with simple structures and with a strong foot- wall and hangingwall. In recent years, as more complex folding and flatter dipping orebodies have been encountered, considerable variations have been introduced. Stope drilling is effected with drifters and extension steel.

Jan 1, 1981

By M. K. Yen, W. R. Hibbard

Previous studies of plastic deformation of metals have emphasized the important role of bending and constraints during strain under relatively pure stresses.1"5 Some new phenomena such as early conjugate slip6 and polygonization7 are intimately concerned with the relief of bending stresses, the former by slip and the latter by a process analogous to re-crystallization. However, few analyses of the bending deformation of single crystals were found in the literature, and those8,9 are sufficiently previous to the modern interpretations of slip to invite further investigation in this area. The transverse bending mode of testing, using the two-point loading method, was selected because the center portion of the specimen may be considered as under a theoretically pure bending stress with a uniform bending moment. Considerable attention was given to both the lattice distortion and flow characteristics of the crystals during deformation. Experimental Proeedare A specimen under the action of equal and opposite couples at both ends is said to undergo pure bending. The shear and moment diagrams of a specimen under the two-point loading method of transverse bending are shown in Fig 1. The moment over the middle-third of the span is equal to PL/6, and the vertical shear is zero. Fig 2 (A) illustrates the stress distribution in the elastic range on a plane across the middle of the specimen. It can be seen that the normal stress has a maximum value at the extreme surfaces on the tension and compression sides, and decreases proportionately to zero as it approaches the neutral plane. However, when the plastic range is reached, the stress distribution is no longer a straight line relation, but changes in a manner which can be best illustrated as shown in Fig 2(B). The following calculation and stress analysis are mainly in accordance with the works reported by Timoshenkol0 and Kochendörfer.9 From the simple beam formula, the maximum normal stress will occur at the surface of the specimen. For a round specimen with a radius r and for a rectangular specimen with a height 2h and width 2b, the maximum normal stress, Sp, will be given by the following formulas: For round specimen: 2PL Sp = 3pr3 [1] For rectangular specimen: Sn = 8bh2 [2] Since it is reasonable to consider that slip takes place under the same conditions as in uniaxial loading, the resolved shear stress S?, along the operative slip direction will be as follows: For round specimen: 2PL sin x cos ? Sp = 3pr3 sin x cos ? [3] For rectangular specimen: PL S8 = 8bh2 sin x cos ? [4] where X is the angle between the specimen axis and the slip direction and X, the angle between the specimen axis and the major axis of the glide ellipse. It was also reported by Kochenclorfer9 that the critical bending moment, that is, the bending moment exerted on the specimen to initiate slip, should be higher than the value given by the equations stated above. Based

Jan 1, 1950

By J. H. Rushton, L. H. Mahony

Principles of fluid motion and turbulence which have been found to be of use in mixing and agitation problems are discussed, as well as suggested applications in extractive-metallurgy processes. Various types of impellers are described, together with other conditions that affect flow pattern and turbulence. The choice of equipment for particular requirements is considered, and equations for power input are given. Modern heavy-duty mixing and agitation equipment can take an increasingly important part in such extractive-metallurgy processes as solids handling, crystallizing and leaching, chemical operations, and flotation. Application of mixing and fluid-mechanics principles to extraction methods can lead to greater process rates and a resultant saving in time and money. MIXERS are being applied with increasing frequency to problems in the metallurgical industries. The increase represents in part the modernization of mechanical equipment used through the application of unitized drives, modern electric motors and speed reducers, and recently developed materials of construction. Some applications have resulted from process changes and the use of techniques for operations similar to those that have been demonstrated and proven in the chemical industry. Still more applications are being developed through the use of fluid-dynamics principles not previously recognized as applicable to these operations. It is the purpose of this paper to describe those principles of fluid motion and turbulence which have been found to be of use in mixing and agitation problems, and to suggest applications for them in the extractive-metallurgy processes. Modern mixing equipment has been designed to provide the fluid motion consistent with that best suited to different operations.' When properly applied, these more efficient fluid-handling mixers result in lower process-production costs. Metallurgical operations usually involve large amounts of solids. The characteristics of the ore control the process and only enough liquid is added to provide sufficient fluidity for handling and processing. The reactive part of the ore, the mineral value to be extracted, is likely to be low in percentage, and the quantity of reagent required is similarly low, but excessive dilution makes additional reagent necessary and therefore is undesirable. Thus, the primary problems in minerals processing are to handle as high a concentration of solids as possible and to distribute reagents uniformly. These pulps, or slurries, are in a range of solids concentration where hindered settling rates apply, and care must be taken to agitate and mix sufficiently to maintain suspension and flow throughout the equipment. Such slurries have high densities which are the weighted average of the components and viscosities that are spoken of as "effective viscosities" but are difficult to evaluate. It is convenient to separate metallurgical operations involving mixers and agitators into three categories: solids handling, mechanical or physical operations, and chemical operations. In each of these categories fluid motion is required to maintain suspension, to distribute surface-active agents, to blend components and reactants, and to induce high rates of chemical reaction. After a discussion of some basic principles of mixing and agitation, examples of application to the three categories will follow. Fluid Motion The fundamental problems of mixing and agitation of liquids have to do with the mechanics of fluid streams and the means by which they are

Jan 1, 1955

By S. S. Solomons, V. H. Tillmann, L. Smaniotto

INTRODUCTION The Zinc Corporation, Limited and New Broken Hill Consolidated Limited (ZC/NBHC) mines are situated at Broken Hill, New South Wales, Australia (refer Fig. 1). Both companies are wholly owned subsidiaries of Australian Mining and Smelting Limited (AM&S). Annual production from both mines is 2.3 million tonnes of lead, silver and zinc ore mined by a variety of underground methods. From 1911 to 1968, conventional overhand cut and fill stoping (both open and square set timber) was used almost exclusively, though some trials with shrinkage and sublevel stoping were undertaken. The trial stopes used short holes in conparison with today's sublevel open stopes. Mechanised cut and fill methods with regular rockbolt support were introduced in the early 1960's and undercut and fill stoping replaced square set timber stoping for pillar extraction in the late 1960's. Since 1968, increasing emphasis has been placed on longhole open stoping (LHOS) although some areas of the orebodies are comrnrnitted to the abovemehtioned methods since existing stope development makes it impracticable to change. In 1979, approximately half the total production was mined by LHDS and it is expected that the LHOS proportion will increase in the future. A vertical crater retreat method of longhole mining has also been successfully introduced to replace the undercut and fill stoping of some pillars and it is expected that this stoping method will be used in some future primary ore extraction. This paper reviews the progress of LHOS technology at ZC/NBHC with emphasis on the stope planning and design process and the importance of rock mechanics applications. The objective has been to maximise the primary extraction of the orebody giving due consideration to stability and safety. This has led to the design of a novel sloped wall longhole stoping method which should eliminate the formation of pillars by mining the orebody in a "continuous" manner incorporating a sloped sand interface between the blocks of ore being mined and the previously mined sandfilled stope. The expected results include the complete extract of the orebody without the formation of pillars and with minimal dilution of ore with sandfill. The evolution of the method can be best illustrated by following the mining progress of the relatively new Southern 'B1 Lode orebody at the NBHC mine. GEOLOGY Southern '8' Lode is the southern extension of the 'Bt Lode orebody, which is currently being mined further to the north in ZC/NBHC (refer Fig. 2). The ore block being developed and stoped extends for a vertical height of 90 metres from No. 17 level down to approximately ten metres above No. 19 level. The ore minerals are sphalerite and galena with the zinc:lead ratio averaging approximately 3:l with a minor amount of silver associated with the galena. There is a sharp contact between the ore and the country rock on the footwall and keel, while the hanging wall gradates into medium and low grade material. The gangue is primarily quartz together with garnet, biotite, feldspar, gahnite, sericite and minor apatite. In various horizons within the orebody, grey feldspar, quartz and garnet pegmatites occur as stratiform layers. The ore and country rock are competent with some inconpetent boundaries.

Jan 1, 1981

By James P. Kimball

Sulphur is always present in mineral coal of every variety. In the oxidized state it may exist as sulphuric acid in combination with a base. In the unoxidized state it exists in combination with iron as bisulphide of iron. It is also supposed by some chemists to exist in unknown combinations with the organic elements of the coal. This supposition requires the demonstration as a fact that certain coals yield more sulphur than what may exist in a free state, or what belongs to recognized combinations, not only with the accompanying iron but with other inorganic bases abundantly occurring in coals, but which unfortunately are very rarely estimated. This practically as well as theoretically important question it is at present impossible to decide, simply for want of available thorough ultimate coal analyses including the ash. Such analyses, at the best imperfect, are luborious and intricate, and therefore seldom made, the much simpler proximate analyses, so-called, being resorted to by most chemists for technical purposes, and commonly substituted for more scientific work even in the laboratories of government surveys. And it is unfortunate enough that few ultimate analyses are extended to the composition of the ash, much less to its relation to the original constitution of the coal. Sulphur chiefly occurs in most coals in the form of yellow pyrites (pyrite) and white pyrites (marcasite), both minerals, characterized by differences in crystalline form and in physical properties as well as in color, but having the same chemical composition (FeS2). No coal is entirely free from one or both of these dimorphous forms of bisulphide of iron. It may be disseminated through the mass so as to be invisible, in plates or scales between the laminæ of the coal, or in conspicuous segregations or nodules, sometimes of considerable dimensions. Coals greatly differ as to the proportion of sulphur which they contain. They also differ as to the relative proportion in which its several forms of combination occur. Such qualitative and qnantitative disparities arc scarcely, if at all, more marked in separate coal seams of a uniform type than in different parts of one and the same coal seam, whether it be considered locally as constituted of several

Jan 1, 1880

By James P. Dr. Kimball

SULPHUR is always present in mineral coal of every variety. In the oxidized state it may exist as sulphuric acid in combination with a base. In the unoxidized state it exists in combination with iron as bisulphide of iron. It is also supposed by some chemists to exist in unknown combinations with the organic elements of the coal. This supposition requires the demonstration as a fact that certain coals yield more sulphur than what may exist in a free state, or what belongs to recognized combinations, not only with the accompanying iron but with other inorganic bases abundantly occurring in coals, but which unfortunately are very rarely estimated. This practically as well as theoretically important question it is at present impossible to decide, simply for want of available thorough ultimate coal analyses including the ash. Such analyses, at the best imperfect, are laborious and intricate, and therefore seldom made, the much simpler proximate analyses, so-called, being resorted to by most chemists for technical purposes, and commonly substituted for more scientific work even in the laboratories of government surveys. And it is unfortunate enough that few ultimate analyses are extended to the composition of the ash, much less to its relation to the original constitution of the coal. Sulphur chiefly occurs in most coals in the form of yellow pyrites (pyrite) and white pyrites (marcasite), both minerals, characterized by differences in crystalline form and in physical properties as well as in color, but having the same chemical composition (FeS2). No coal is entirely free from one or both of these dimorphous forms of bisulphide of iron. It may be disseminated through the mass so as to be invisible, in plates or scales between the laminæ of the coal, or in conspicuous segregations or nodules, sometimes of considerable dimensions. Coals greatly differ as to the proportion of sulphur which they contain. They also differ as to the relative proportion in which its several forms of combination occur. Such qualitative and quantitative disparities are scarcely, if at all, more marked in separate coal seams of a uniform type than in different parts of one and the same coal seam, whether it be considered locally as constituted of several

Jan 1, 1880

By Bernt Ellingsaeter, Terkel Rosenqvist

IN the silicothermic reduction of magnesia, burned dolomite is treated with high grade ferrosilicon in an evacuated steel retort at temperatures between 1150° and 1200°C. The following reaction is assumed to occur MgO + CaO + 1/2 Si (Fe) - Mg(g) +- & Ca,SiO, + Fe with the equilibrium constant on the basis of thermodynamic data for the substances involved and assuming unit activity for all solid constituents, Kubaschewski and Evans' calculated the magnesium pressure as a function of temperature logp,, (mmHg) = 15.06-1.61 T- log T- 13,16O/T. Pidgeon and King," who measured the magnesium pressure by means of a flow method, found values about three times the calculated ones. This may be caused by uncertainties in the thermodynamic data, but may also be caused by the formation of other reaction products, such as Ca,SiO,, which is known to be the stable silicate in equilibrium with lime. Whatever the exact reaction may be, it is evident that as long as all other factors are maintained constant the magnesium pressure will vary with the silicon activity, viz: p,, = &,. (p,,) ", where (p,,)" is the magnesium pressure by the reduction with silicon of unit activity. In the present investigation, p,,, was therefore measured for various grades of ferrosilicon and their silicon activities were calculated. The experimental method was similar to that used by Pidgeon and King. Briquettes of burned dolomite and Fe-Si alloys were placed in a 1 in. heat resistant steel tube and were heated to 1200°C. By means of a flow of purified hydrogen, the magnesium vapor was swept into a condenser filled with steel wool. From the amounts of hydrogen used and magnesium condensed the magnesium partial pressure was calculated. First a tube of 28 pct Cr steel, and later a tube of Kanthal D (a ferrous alloy with 22.6 pct Cr, 2 pct Co, and 4.5 pct Al), was used as a heat resistant tube. The latter tube showed the best heat resistance. The dolomite used had the following composition: 43.6 pct MgO, 55.5 pct CaO, 0.2 pct K,O + NalO, 0.2 pct Fe,O,, and 0.3 pct insoluble. The ferrosilicons were prepared from 98 pct Si of technical grade (0.7 pct Al, 0.3 pct Ca, and 1 pct Fe) by melting with various amounts of electrolytic iron of high purity. The alloys thus obtained were analyzed on their iron and silicon contents, which added up to between 99 and 100 pct. The amounts of reactants were chosen so that there would always be an excess of the oxides, and the average decrease in silicon content during the experiments would be about 1 pct. Most of the reaction occurred in the part of the reaction mixture facing the gas flow. The change in the silicon concentration in the part facing the condenser was, therefore, negligible. To facilitate the reaction, about 1 pct fluorspar was added to the briquette mixture. Fluorspar does not enter the reaction, but is known to act as a

Jan 1, 1957

By W. E. Newell

The basic principles and assumptions involved in D. L. White's solution5 for ultrasonic wave amplification in piezoelectric semiconductors are summarized. If the gain per unit length is maximized at each frequency without regard to the drift-field power density, the gain increases linearly with frequency. Therefore it would seem that very high gain is possible for frequencies in the hundreds of megacycles. However, with present materials the resulting drift-field power density is so large that cooling would be a major problem. For this reason it is more realistic to judge the capabilities of an ultrasonic amplifier in terms of the gain that is possible with a specified power density. In this case the gain at high frequencies does not increase but decreases more rapidly than l/f. The calculation of the gain us frequency at a fixed power level involves the repeated solution of a cubic equation and is therefore tedious. For the purpose of comparing different materials, sufficient accuracy can be obtained from an approximate solution which may easily be plotted as three straight lines on logarithmic graph paper. This method is used to compare the Performances of CdS, CdSe, ZnO, and GaAs. At a power density of 10 watts per cu cm, the first three give maximum gains of 60 to 80 db per cm in the vicinity of 60 megacycles. At the same power density, GaAs has a maximum gain of about 7 db per cm near 150 megacycles. Although an increase in mobility usually increases the gain for a specified power density, it is shown that there is an optimum mobility above which the gain decreases. An approximate expression for the optimum mobility is established. It is well-known that the interaction between charged particles and an electromagnetic wave can be used to amplify the electromagnetic wave. In the conventional traveling-wave tube, electrons in a vacuum are caused to move at a velocity slightly greater than the wave velocity, thereby coupling energy into the wave.' In principle, the same effect would be evident in a solid except that repeated collisions prevent the electrons from attaining a velocity greater than that of an electromagnetic wave. However, the drift velocity of electrons can be made to exceed the velocity of an acoustic wave in a solid. In 1956, weinreich2 noted that the deformation potential provided one means of coupling energy from drifting electrons into an acoustic wave, but the effect was too small to be of any practical importance. It was not until 1961 that White and Hutson demonstrated that the piezoelectric effect provided a strong enough coupling to give appreciable amplification of an ultrasonic wave and awakened wide interest in electron-phonon interactions. Qualitatively the effect may be described as follows. An acoustic wave propagating through a piezoelectric material is accompanied by an electric-field wave. If the vibrations are properly oriented with respect to the crystal axes of a piezoelectric semiconductor, the local electric fields are axially directed and cause bunching of the charged carriers. If the drift velocity of these carriers is less than the velocity of sound, they tend to be dragged along by the acoustic wave. This electroacoustic effect results in a net potential appearing across the material and in attenuation of the acoustic wave. On the other hand, if the drift velocity exceeds the velocity of sound, energy is coupled from the carriers to the acoustic wave and amplification results. Related effects which are being investigated include ultrasonic amplification in semimetals6-' and semi-conductor, and current saturation in piezoelectric semiconductors.'4-17 There are various problems which hinder the immediate application of ultrasonic amplification in a practical device. First, there are the traditional problems encountered in the design of acoustic delay lines: attenuation in the electromechanical transducers, reflected waves (which can lead to self-oscillation in an amplifying device), mechanical tolerances necessary to maintain phase coherence, and so forth. Second, there is the problem of selecting the best material from among the various piezoelectric semiconductors as they become available in single-crystal form with a wide choice of characteristics. This paper deals with the second problem and shows that the most realistic way of comparing materials is in terms of the gain which can be obtained without exceeding a

Jan 1, 1964

By B. G. Zambre, G. A. Parks, J. N. Roco

Automatic controls have long been a fact of technological life. Some understanding of control systems is rapidly becoming essential to the competence of mineral engineers. Students in undergraduate mineral engineering curricula, however, should not be expected to develop proficiency in the design and interpretation of control systems. They should be provided the opportunity to absorb some of the fundamental concepts and cant of such highly specialized peripheral fields as automatic controls. The present paper describes the components of a simple, flexible and relatively inexpensive control system designed primarily for the use of students in independent study projects. The control system is based on a continuous on-stream optical analyzer analogous to commercially available systems using X-Ray fluorescence analysis. The components can be easily adapted to control other separators and separations. As described here, they have been arranged to control the magnetic separation of a magnetite bearing specular hematite from quartz to produce a quartz product of fixed hematite content using a minimum of power. The control system consists of the magnetic separator* with a vibrating feeder, † a sampler, an optical scanner, a differential amplifier, a millivolt recorder ‡ and a motor driven, metered variable DC power supply for the separator electromagnet. An optional magnet current recorder is useful. The system as arranged for closed loop control is illustrated in Fig. 1. The elements of the loop are identified in automatic control jargon in Table I. The same components can be easily adapted for use in other control modes.§.1 * A Carpco Induced Roll High Intensity Magnetic Separator (Carpco Manufacturing, Inc., Jacksonville, Fla.); modified for external magnet power supply. A Jeffrey-Traylor Vibrating Feeder Type 1B (Jeffrey Manufacturing Co., Columbus, 0.). A Beckman Laboratory Potentiometric Recorder (Beckman Instruments, Inc., 619-A) fitted with adjustable limit switches for external circuit control (Beckman Instruments, Inc.. 71980). Terminology is taken from ref. 1. CLOSED LOOP OPERATION The separation is monitored through the optical reflectivity of the tailings stream which is proportional to the hematite content, THA. The difference, A, between the measured analysis (actually a proportional voltage) and a reference analysis (also a proportional voltage) is continuously recorded. The separation is controlled through changes in the electromagnet current, I, supplied from a motor driven variable source. The two primary variables, ? and I, are linked through adjustable limit switches on the

Jan 1, 1965

By Grover J. Holt

In order to provide a clear picture of the development of the sink-and-float process of concentration as applied to the iron ores of Minnesota, a few pertinent facts should be brought out concerning the ores and the products of concentration. Iron ore is primarily a cheap commodity handled in immense quantities. The price of iron ore at present, based on a natural iron content of 51.5 per cent, varies from $2.55 to $2.70 per gross ton f.0.b. cars at the mine or washing plant in Minnesota. Last year 92,000,000 gross tons were shipped from the Lake Superior district (Minnesota, Michigan, and Wisconsin) to the furnaces. Concentrated iron ores have no greater value per unit of iron content than ores that can be mined and shipped direct from the shafts or open pits. For years only the higher grades of concentrating ores could be treated and sold in competition with the immense reserves of direct shipping ore available. These higher-grade "straight wash" ores required only the simplest treatment; namely, the washing out of fine sand from coarse ore particles. As the exhaustion of the " straight wash " ore is reached, rock is found replacing part of this sand, and the ore no longer responds to the former sizing treatment. Substantial tonnages of this marginal (jig ore) remained to be treated. Because of the limitations placed on concentrates by the furnaces—namely, low price of product and minimum amount of fines—iron-ore concentration oil has been and must continue to be a one-pass process. Only an increase in price or a change in structure of concentrates allowed by the ore consumer will encourage departure from this practice. Faced with the necessity of handling large tonnages, with the limitations mentioned, iron-ore research men have found their search for new and more efficient methods of treatment seriously handicapped and removed from the field of fine crushing. Sink-and-float treatment appeared to offer the best field in coarse concentration and has been proved by several plants to be well adapted to our problem. The use of ferrosilicon as a medium solid was decided upon. In our present practice, the ferrosilicon metal, carrying about 83 per cent iron, 15 per cent silicon and 2 per cent carbon and nonmagnetics, is crushed to pass 65 mesh and used without further sizing. When mixed with water the specific gravity of the medium may be controlled to within 0.02 per cent above or below the required gravity. Table I shows examples of a change in solid content of the medium for varying gravities. the I is based on clean ferrosilicon, whereas some contamination is always present in the mill circuit, tending to

Jan 1, 1947

By Grover J. Holt

In order to provide a clear picture of the development of the sink-and-float process of concentration as applied to the iron ores of Minnesota, a few pertinent facts should be brought out concerning the ores and the products of concentration. Iron ore is primarily a cheap commodity handled in immense quantities. The price of iron ore at present, based on a natural iron content of 51.5 per cent, varies from $2.55 to $2.70 per gross ton f.0.b. cars at the mine or washing plant in Minnesota. Last year 92,000,000 gross tons were shipped from the Lake Superior district (Minnesota, Michigan, and Wisconsin) to the furnaces. Concentrated iron ores have no greater value per unit of iron content than ores that can be mined and shipped direct from the shafts or open pits. For years only the higher grades of concentrating ores could be treated and sold in competition with the immense reserves of direct shipping ore available. These higher-grade "straight wash" ores required only the simplest treatment; namely, the washing out of fine sand from coarse ore particles. As the exhaustion of the " straight wash " ore is reached, rock is found replacing part of this sand, and the ore no longer responds to the former sizing treatment. Substantial tonnages of this marginal (jig ore) remained to be treated. Because of the limitations placed on concentrates by the furnaces—namely, low price of product and minimum amount of fines—iron-ore concentration oil has been and must continue to be a one-pass process. Only an increase in price or a change in structure of concentrates allowed by the ore consumer will encourage departure from this practice. Faced with the necessity of handling large tonnages, with the limitations mentioned, iron-ore research men have found their search for new and more efficient methods of treatment seriously handicapped and removed from the field of fine crushing. Sink-and-float treatment appeared to offer the best field in coarse concentration and has been proved by several plants to be well adapted to our problem. The use of ferrosilicon as a medium solid was decided upon. In our present practice, the ferrosilicon metal, carrying about 83 per cent iron, 15 per cent silicon and 2 per cent carbon and nonmagnetics, is crushed to pass 65 mesh and used without further sizing. When mixed with water the specific gravity of the medium may be controlled to within 0.02 per cent above or below the required gravity. Table I shows examples of a change in solid content of the medium for varying gravities. the I is based on clean ferrosilicon, whereas some contamination is always present in the mill circuit, tending to

Jan 1, 1947

By Renato G. Bautista, Robert A. Hard

A comparative study has been made of the ex-tractability of several of the transitim metals from thiocyanate sohtions using methyl isobutyl ketone as the organic solvent. Extractions were made of scandium (III), chromium (III), manganese (II), iron (III), and cobalt (II), of which only the scandium , iron, and cobalt were readily extracted. All extractions were made at pH 2 and at 25.5 °C. A series of extraction experiments was made on scandium, iron, and cobalt to determine the effect of thiocyanate cmcentration added as the ammonium and also as the calcium salt. Also determined was the effect of metal ion concentratim and ionic strength on the distribution ratio. Analyses of the organic extracts indicated the scandium and iron were extracted principally in the uncharged trithiocyanate form, while the extractable form of the cobalt contained the cobaltotetrathiocyanate anion. THE chemical separation of metal complexes in solution by solvent extraction has in recent years been developed for large scale separation as well as bench scale work in the laboratory. It has also found increased application in analytical chemistry. Its extension to extractive metallurgy as a unit operation is best illustrated in the recovery of uranium and vanadium from their leached liquor.5-8 The extractability of some thiocyanate systems have been extensively investigated. The rate earths have been separated from one another using butanol as a olvent. Scandium was effectively extracted in ethyl ether using ammonium thiocyanate as the com-plexing agent.'' Similarly, the separation of hafnium from zirconium has been reported using ether""' or hexone13 as the organic solvent. The extractability of various metal thiocyanates in ethyl ether is given by 0ck.l Sharp and Wilkinson15 prepared nickel-free cobalt slats by thiocyanate extraction. The mechanism of extraction of titanium (IV) thiocyanate with methyl isobutyl ketone was determined by elafosse.''" The characterization of cobalt (TI) thiocyanate complex using methyl isobutyl ketone as the solvent at near neutral aqueous solution and the spectrometric studies of the ex-tractable cobalt complex has been made by Hard." It is noteworthy to mention at this point that the fundamental characteristic of ketone (oxonium) solvent extraction is that the extractable species are uncharged ion pairs solvated by the ketone molecules, but probably still partially hydrated.

Jan 1, 1963

By John McDonald

THE amount of published literature in the field of mechanical proper-ties of magnesium alloys is not great; particularly with respect to rolled alloys. Haughton and Prytherch1 have summarized most of the articles to 1937, including their own work on sheet. Although many papers can be found dealing' with the technology of magnesium and the alloys now in commercial use, there does not seem to be any work dealing exclusively with the variation in properties of rolled alloys as caused by changes in alloy composition. Comparison of published data reveals considerable discrepancies in, properties of alloys of the same composition, probably due to differences in the details of working and heat-treating followed by different workers. The purpose of this paper is to study the proper-ties of alloys in conditions where the effect of such differences have been eliminated. It is well known that the strength and ductility of metals are functions of the working and heat-treating they have received, as well as of their chemical composition. In comparing alloys with respect to these proper-ties, therefore, all possible combinations of the variables of the fabrication procedure ought to be carried out on each alloy. One or more of these treatments will produce the highest ductility, or strength, or both. These highest properties would then be selected as characteristic of each alloy, and alloys would be compared on the basis of those properties. Thus a separation of the effect of composition and that of working and heat-treating would be achieved. In this paper, however, a compromise with completeness has been made by carrying out a restricted number of treat-ments. As a consequence, the properties obtained are not necessarily the best that the alloy is capable of exhibiting; but since all alloys have been treated by the same method, the results ought to be comparable. The elements selected for the study comprise a class that will alloy easily with magnesium, and that are relatively cheap. Each was added in different amounts, up to a limit that varied with the element.

Jan 1, 1939

By Hugh D. Pallister, Richard W. Smith

The Alabama flake-graphite industry has flourished only in times of war when importations of foreign graphite for crucible use have been greatly curtailed or cut off. World War I was a boom period and in 1918 some 30 plants produced 7,795,475 Ib. of flake graphite. The industry languished following the war and no graphite was produced from 1929 to 1939. The Federal Bureau of Mines began investigating the Alabama graphite deposits in 1940 and worked out modern recovery methods that were applied to production during part of World War 11. A prospecting program disclosed reserves of over 25 million tons of graphtic schists. In 1942 the War Production Board authorized the enlargement of the only plant then producing flake graphite and the construction of two new plants. Production of flake graphite from these three plants in 1943 is estimated at 8,100,000 lb. Increased imports from Madagascar late in 1943 shut down one of the new plants and curtailed production at the other. Alabama flake graphite has not yet been able to compete with Madagascar graphite for use in the manufacture of any but small graphite crucibles. The future of the industry must depend on meeting the higher carbon content demanded for other uses. Introduction The flake-graphite deposits of Alabama are confined to a narrow belt extending for about 60 miles southwest from the northeast corner of Clay County across Coosa County to the edge of the Coastal Plain in Chilton County (Fig. I). The continuity of the beds is broken by a 10-mile gap between Millerville in Clay County and Goodwater in Coosa County. The Clay County beds extend about 27 miles northwest of this gap, with a width of 3 to 4 miles near Millerville and of less than a mile at the northeast end. The Coosa-Chilton are? is about 33 miles long with a fairly consistent width of 2 to 3 miles. Geologically the graphite belt lies near the northeast boundary of the outcrop of metamorphic rocks of the Ashland series, a complex group of intensely folded and faulted beds of pre-Cambrian age composed mostly of quartz-mica schist, in which the mica is predominantly muscovite; garnet-mica schist, in which biotite is more common; and hornblende schist. Flake graphite is found in commercial quantities only in the quartz-mica schist. The best graphite, both in quantity and quality, is found in schist cut by numerous small pegmatite stringers along the planes of the schistosity. The pegmatites themselves contain little graphite, but the adjacent country rock is normally rich. . The schist was derived through the recrystallization of shale during the intrusion of the pegmatites, the graphite being derived from amorphous carbon in the shale. The intrusion of the pegmatites, and therefore the size and quantity .of the graphite flakes, seems to have a structural control.

Jan 1, 1948

By Hugh D. Pallister, Richard W. Smith

The Alabama flake-graphite industry has flourished only in times of war when importations of foreign graphite for crucible use have been greatly curtailed or cut off. World War I was a boom period and in 1918 some 30 plants produced 7,795,475 Ib. of flake graphite. The industry languished following the war and no graphite was produced from 1929 to 1939. The Federal Bureau of Mines began investigating the Alabama graphite deposits in 1940 and worked out modern recovery methods that were applied to production during part of World War 11. A prospecting program disclosed reserves of over 25 million tons of graphtic schists. In 1942 the War Production Board authorized the enlargement of the only plant then producing flake graphite and the construction of two new plants. Production of flake graphite from these three plants in 1943 is estimated at 8,100,000 lb. Increased imports from Madagascar late in 1943 shut down one of the new plants and curtailed production at the other. Alabama flake graphite has not yet been able to compete with Madagascar graphite for use in the manufacture of any but small graphite crucibles. The future of the industry must depend on meeting the higher carbon content demanded for other uses. Introduction The flake-graphite deposits of Alabama are confined to a narrow belt extending for about 60 miles southwest from the northeast corner of Clay County across Coosa County to the edge of the Coastal Plain in Chilton County (Fig. I). The continuity of the beds is broken by a 10-mile gap between Millerville in Clay County and Goodwater in Coosa County. The Clay County beds extend about 27 miles northwest of this gap, with a width of 3 to 4 miles near Millerville and of less than a mile at the northeast end. The Coosa-Chilton are? is about 33 miles long with a fairly consistent width of 2 to 3 miles. Geologically the graphite belt lies near the northeast boundary of the outcrop of metamorphic rocks of the Ashland series, a complex group of intensely folded and faulted beds of pre-Cambrian age composed mostly of quartz-mica schist, in which the mica is predominantly muscovite; garnet-mica schist, in which biotite is more common; and hornblende schist. Flake graphite is found in commercial quantities only in the quartz-mica schist. The best graphite, both in quantity and quality, is found in schist cut by numerous small pegmatite stringers along the planes of the schistosity. The pegmatites themselves contain little graphite, but the adjacent country rock is normally rich. . The schist was derived through the recrystallization of shale during the intrusion of the pegmatites, the graphite being derived from amorphous carbon in the shale. The intrusion of the pegmatites, and therefore the size and quantity .of the graphite flakes, seems to have a structural control.

Jan 1, 1948

By Strathmore R. B. Cooke, Marcus Digre

That calcium will activate quartz for flotation with anionic collectors such as soaps has been known for a number of years,12'3 and the method has been applied to the concentration of various iron ores4 5 6 as well as to other materials. Clemmer gives the permissible pH range for the activation of quartz by calcium as between 9.5 and 12.5 with an optimum value of 11.0, and Hertzog7 has determined the calcium abstraction of iron ores and of Ottawa sand, but otherwise little quantitative data are available. The purpose of the present investigation was to determine those conditions under which dilute solutions of calcium chloride will activate quartz for anionic flotation and to determine the amount of calcium ion adsorbed by quartz. Those variables, the influence of which was primary, such as calcium, sodium, and hydrogen ion concentrations, were studied. Other factors which influenced the results are discussed in the description of the experiments. Three different experimental methods were used to determine the activation of the quartz. The first method was to treat a sample of ground, deslimed quartz with solutions of calcium chloride and sodium hydroxide, and to determine the calcium adsorbed by chemical analysis of the solution. Advantages are that a direct value for the amount of calcium adsorbed is obtained, and introduction of a collector which may affect the adsorption is avoided. The difficulty is to determine the extremely small differences in the calcium concentrations before and after the abstraction. The second method was to determine the activation of the quartz by observing the contact angle obtained on pressing an air bubble against the polished surface of a piece of quartz immersed in a solution of calcium chloride, sodium hydroxide, and sodium oleate. This is the well-known "contact angle" method. The third method was to determine those conditions under which an air bubble will pick up quartz particles immersed in a solution of the desired reagents. Disadvantages common to the sec- ond and third methods are that introduction of a collector is necessary to obtain bubble contact with the mineral; very small quantities of contaminant can give erroneous results; and no information is obtainable concerning the amount of calcium adsorbed. Advantages are that tedious analyses, which at best give approximate results, are avoided, and the conditions, especially for the third method, closely simulate those of actual flotation. Although the most consistent results were obtained with the third method, this paper describes the techniques used and the results obtained from each of the three methods. Distilled water and analytical grade reagents were used throughout except where otherwise stated. Between tests all glass articles were cleaned with a chromic-sulphuric acid mixture and sodium hydroxide. Abstraction Method The raw material for the tests was Ottawa sand sized between 14 and 28 mesh. Batches of 500 g of this sand were dry ground for 1 hr in an Abbe porcelain mill with 2700 g of quartz pebbles. The batches were split down to samples weighing approximately 30 g, and the weight of each sample was adjusted to exactly this value.

Jan 1, 1950

By S. H. Ash

Diesel locomotives, trucks, bulldozers, and other diesel-powered equipment are fast proving their superiority for mine transportation purposes. MORE than 100 minerals are mined and processed in the U. S. Management and labor have negotiated wage-scale agreements that have balanced wages and affected cost of labor in such a manner that comparisons can be made of transportation costs. Transportation of materials underground has become of vital importance.1"0 This report describes underground haulage methods and equipment, presents cost analyses, and discusses future plans for underground transportation in metal mines. It does not cover hoisting in shafts or slopes by cage, skip, or cars. Advancements have been made by the metal mining industry in lowering transportation costs in several mines, making it possible to mine orebodies or portions of orebodies not considered economically recoverable a few years ago.7,8 Wumerous reports have been written on transportation at metal mines; this report covers recent haulage at some representative mines throughout the U. S. The term metal mine is used to include all noncoal mines from which up-to-date information was obtainable. The best haulage equipment is subject to break- downs and accidents; but unless the rolling equipment, track, and signal systems are kept in good repair and are operated systematically, these breakdowns and accidents are likely to be frequent, destructive, and more costly than are necessary.10 Rail haulage crews using dependable rolling stock can move trips safely over grades low enough to be traveled without loss of control; this applies particularly if the track is straight and well maintained. Few accidents occur on main haulage lines if visibility and clearance are good and clear track is assured. One-way traffic underground is a recent development contributing to reduction of accidents and increased efficiency in mine haulage (Kelly mine, Anaconda Co., Butte, Mont., and Climax mine, Climax Molybdenum Co., Climax, Colo.)." Locomotives: Six types of mine locomotive used in recent years are trolley, storage battery, combination trolley and storage battery, diesel, compressed air, and gasoline. Electric haulage has been displaced by compressed air locomotives, which have been used successfully in certain instances, especially on dirty and wet tracks and steep grades." Maintenance cost is high, and operating efficiency, from a power standpoint, is low. Compressed air locomotives operate under self-contained power, and the absence of bare elec-

Jan 1, 1957

By J. Murray Riddell, Grover J. Holt, Arthur E. Anderson

Out of the depths of each business cycle we emerge with a stimulus for greater efficiency and a realization of progress in industrial technique. The recent years have not been an exception to this rule, and in retrospect we sense that significant advances have been made in mining practices on the Mesabi Range, with emphasis on speed of production and economy of operation. A survey of such changes does not reveal anything that is fundamentally or basically new, but it is apparent that ideas and machines that have found wide application in industry elsewhere, with necessary modifications, are being put to productive use in mining iron ore. Innovations in transporting and conveying ore from the open pits stand out with greatest prominence, so that, in the main, it will be with these features that this paper will treat, in the review of progress made to this date. The late years have witnessed the development of small deposits for open-pit operation, the cleaning up of pits nearing exhaustion, and in two large operations the approach to the limit for haulage by steam locomotives, owing to grades, making it necessary to recover the ore in slopes, track benches and below the power-shovel limit by a milling and scram operation. For several years small standard-type dump trucks have been used in clean-up or scram operations within the pits, but during the season of 1936 the first effort was made to use heavy truck haulage in the transportation of stripping and iron ore. The new units in use today vary from 15 to 20 tons in capacity, and are equipped with powerful gasoline or Diesel engines and supplemented by trailers or wagons on rubber or crawler treads. The Diesel engine finds favor with most operators because of fuel economy. These units develop upward of 125 hp. in the 15-ton trucks. The speed of these trucks varies from 2 miles per hour in Pow gear to 35 in high gear. The transmission affords 12 speeds forward and 3 speeds in reverse.

Jan 1, 1940

By J. Murray Riddell, Arthur E. Anderson, Grover J. Holt

Out of the depths of each business cycle we emerge with a stimulus for greater efficiency and a realization of progress in industrial technique. The recent years have not been an exception to this rule, and in retrospect we sense that significant advances have been made in mining practices on the Mesabi Range, with emphasis on speed of production and economy of operation. A survey of such changes does not reveal anything that is fundamentally or basically new, but it is apparent that ideas and machines that have found wide application in industry elsewhere, with necessary modifications, are being put to productive use in mining iron ore. Innovations in transporting and conveying ore from the open pits stand out with greatest prominence, so that, in the main, it will be with these features that this paper will treat, in the review of progress made to this date. The late years have witnessed the development of small deposits for open-pit operation, the cleaning up of pits nearing exhaustion, and in two large operations the approach to the limit for haulage by steam locomotives, owing to grades, making it necessary to recover the ore in slopes, track benches and below the power-shovel limit by a milling and scram operation. For several years small standard-type dump trucks have been used in clean-up or scram operations within the pits, but during the season of 1936 the first effort was made to use heavy truck haulage in the transportation of stripping and iron ore. The new units in use today vary from 15 to 20 tons in capacity, and are equipped with powerful gasoline or Diesel engines and supplemented by trailers or wagons on rubber or crawler treads. The Diesel engine finds favor with most operators because of fuel economy. These units develop upward of 125 hp. in the 15-ton trucks. The speed of these trucks varies from 2 miles per hour in Pow gear to 35 in high gear. The transmission affords 12 speeds forward and 3 speeds in reverse.

Jan 1, 1940