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By I Ritchie, G Lane, C Morley, D Yataco

Marcobre SAC (Marcobre) is developing the Mina Justa project in Peru. The project has a lower grade oxide orebody overlaying the higher grade sulfide orebody. The oxide ore is treated by acid leaching; consumption of acid is high and is managed by vat leaching rather than the more conventional heap leaching process.The leaching process uses ore crushed to 100% passing 8 mm to optimise copper recovery. Acid consumption generally increases with a higher proportion of fines. Marcobre is investigating a three-stage crushing circuit with high pressure grinding rolls (HPGR) in the third stage, instead of the more conventional four-stage cone crushing circuit. The use of HPGR can result in higher fines production due to the higher compression forces in the crushing zone. This paper reports on the outcomes of the test work completed by Marcobre in 2016 and 2017 to manage fines production and compare the use of HPGR with conventional crushing on a high acid consuming ore.At moderate pressing forces, a HPGR crushing circuit appeared to produce similar size distributions to cone crushing. On average, both the copper recovery and the acid consumption in subsequent vat leaching were higher with HPGR crushing than with cone crushing.CITATION:Lane, G, Morley, C, Yataco, D and Ritchie, I, 2017. Are there differences in size distribution when comparing conventional crushing and HPGR?, in Proceedings MetPlant 2017, pp 60–72 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Sep 11, 2017

By Thomas Pichler

In the past marine mining technology has focused on bulk mining of base metals from polymetallic massive sulfides. This has not been very successful because of the relatively low unit value of these metals and other reasons such as water depth, etc.. The process of gold enrichment in hydrothermal deposits is controlled through five different factors: (1) nature and source of the solution, (2) source of the gold, (3) a mechanism for mobilizing and transporting, (4) a mechanism for deposition, and (5) a mechanism to preserve the deposit. Gold has been identified in marine polymetallic massive sulfides from various locations on the ocean floor (e.g., spreading centers, seamounts) with contents ranging up to 5 ppm Au. This is within the current range of 1.0 to 7.0 ppm gold content which classifies gold ores. The gold appears to have been derived during hydrothermal alteration of the oceanic crust. Until the present there has been little attention to the possibility of mining gold or other precious metals in the marine environment. However, the magnitude of hydrothermal alteration (seawater can penetrates to a depth of 5 km into the oceanic crust) and primary gold contents of the rocks in the pathway of the fluid indicates that major gold deposits must have formed on or in the ocean floor, but they have not been found. To find a big deposit might only take an exploration model exclusively designed for gold exploration on the ocean floor.

Jan 1, 1993

By W. J. Schlitt

Introduction A major uncertainty in assessing the potential for heap-and dump-leach projects is how to determine the extraction-rate curve for the recovery of the mineral values from coarse ore. Such material could either be run-of-mine (ROM) or primary crushed ore. The problem with field testing coarse ore, especially for new projects, is the large scale and extended leach times needed to accurately determine the final extraction-rate curve. At least 5 x 103 to 5 x 104 t of representative ROM ore are typically required for a copper test heap, and much more is often used. Kennecott, for example, recently constructed a 0.9 Mt (1 million st) ROM test heap at the Bingham Canyon Mine in Utah. In such coarse ore operations, the ultimate level of extraction will require a leach cycle that can extend from several months to a few years. Quite often, project development schedules do not provide the luxury of mining such large quantities of material or running such long tests. Instead, test data are usually limited to results from column leach studies on relatively fine ore, often with a top size that does not exceed 25 mm. Maximum leach times are also short, typically less than a year before an initial decision is needed on project viability. Proposed method One approach to estimating the recovery from a coarse ore leach is to assume that the leach solution will have some ultimate penetration distance into the rock. Then, the final level of mineral extraction in this "leached rim" will equal the ultimate level of extraction identified in various testing programs. Obviously, if the radius of a given rock fragment is less than the penetration distance, that fragment will be fully leached at the end of the operation. With larger rock sizes, the percent recovery will fall off as the size increases and the fraction of unpenetrated rock mass increases. Such an approach sounds simple but is likely to be complex when applied to a real project. For example, the penetration distance will be a function of both the rock characteristics and the effective length of the leach cycle. The important rock characteristics include rock porosity, the degree of internal fracturing and the mode of mineral occurrence. With regard to the latter, penetration is likely to be greater if the leachable mineralization occurs on fracture surfaces or in veinlets, as opposed to fine grains uniformly disseminated throughout the rock mass. An estimate of penetration distance may be derived from column or heap tests by noting the depth of solution penetration into the larger rock fragments after three, six and 12 months of leaching. While the penetration rate is ore specific, something on the order of 10 to 20 mm/y may be appropriate for competent, primary copper (chalcopyrite) ore. For gold in tight quartz, the rate may be about the same. Copper oxide ores and gold that is hosted in a more porous rock matrix are likely to have penetration rates that are at least two to three times higher, and an even higher rate should be appropriate for uranium hosted in sandstone. As noted above, the length of the effective leach cycle is likely to be measured in years. On this basis, the ultimate penetration distance (dp) would vary from less than 50 to several 100 mm when a particular ore is leached to exhaustion. Several sets of mathematical manipulations are necessary to convert a rock size distribution and corresponding value of dp into an estimated extraction-rate curve. The first step is to break the ROM size distribution down into intervals and then calculate the radius for the mean rock size in each interval. This is shown in Table 1 for rock sizes up to 1.75 m (about 6 ft) in diameter. The next step is to calculate the volume of unleached core and the fraction of rock that is leached. This is done for the following three values of dp: 25, 100 and 250 mm. Results are shown in Table 2. The third step is to select the ultimate level of recovery that will be achieved in the fraction of material that is effectively leached, i.e., the outer zone that is penetrated by the leach solution. This is clearly a site-specific factor that can only come from metallurgical test results on representative ore

Jan 1, 1998

By George Simkovich

It was hypothesized that solid-state point defects should alter the flotation properties of solids. Tests conducted on pure AgCl and AgCl doped with CdC12 show that atomic point defects exhibit an important role in the floatability of AgC1. Tests conducted on PbS doped with Ag2s or Bi2S3, also show that the defect structures resulting from these dope additions, i.e., a combination of electronic and atomic point defects, contribute significantly to the flotation of PbS. IT has been established that flotation occurs only when a finite contact angle exists between a solid and a gaseous bubble.' This angle, measured through the liquid phase, is expressed by the equation where the are inter facial free energies and the subscripts S, G, and L represent solid, gas, and liquid phases, respectively. As is seen in Eq. [I] three interface free energies, sG, sl, and GL, enter into the contact angle equation. Therefore, any variation in these energies which sufficiently varies the contact angle will, in turn, vary flotation processes. Changes made in any of the phases concerned, i.e., gas, liquid, or solid phase, are reflected through the changes occurring in two of the surface energy terms. Thus, a change in the liquid composition would be noted in sL and GL, and it is this phase, the liquid, which is most frequently altered in flotation studies., Changes in the solid phase must be reflected through the changes occurring in the sG and sL terms. In particular, it is hypothesized that changes in the surface concentrations of point defects in the solid-phase will alter the sG and sL terms which, in turn, will be reflected by flotation results. As an illustration of this hypothesis one may consider the defect structure and the flotation of AgC1. The bulk defect structure of AgCl is essentially one involving equal number of cation vacancies and interstitial cations.3 Upon adding CdC1, to AgC1, a greater number of silver ion vacancies are created in the bulk of the crystal.4 On the surface of the crystal the smaller binding forces and the free space accomodations may also allow for the creation of "surface interstitial anions", which will be designated as ad-anions. Thus, the point defect structure of the surface of AgCl doped with CdCl, will consist of cation vacancies and/or adanions. If the molecular forces responsible for the surface energies, ?SG and ?sL, are significantly altered by the presence of these surface point defects, then differences in flotation results will be noted as the concentration of these defects is varied. The defects present in AgCl are predominantly atomic in nature. In the case of PbS both electronic and atomic defects are present.5 This compound conducts electrically by either electrons or electron holes depending upon whether excess lead or excess sulfur is present. Upon disolving BiS3 in stoichio-metric PbS, one increases the concentration of cation vacancies and the number of electron carriers in the bulk of the crystal.5" At the surface, the possibility of ad-anions must also be considered. Conversely, upon dissolving AgS in stoichiometric PbS one increases the concentration of interstitial cations and the number of electronhole carriers in the bulk of the crystal.5,6' At the surface the interstitial cations will be designated as ad-cations. Thus, the point defect structure of the surface of a PbS crystal doped with Bi2S3 will consist of a number of cation vacancies and/or ad-anions and an excess of electrons. Conversely the point defects on the surface of a PbS crystal doped with Ag2S will consist of a number of ad-cations and an excess of electron holes. Again, as in the case of AgC1, should the molecular forces responsible for the magnitude of the interface free energies, ?sG and ?sL, be significantly altered by the presence of these surface defects then significant differences in flotation results will be noted as the concentration of these defects is varied. EXPERIMENTAL To test this hypothesis flotation tests were conducted on pure and doped AgCl and on PbS doped with either Bi2S3 or Ag2S. Preparation of the AgCl samples was performed as follows: AgCl and weighed amounts of CdC1, were melted in a porcelain crucible. The melt was then forced through a capillary tube and the particles emitted solidified in air as they fell about 1.5 meters. Spherical particles, -0.50 + 0.25 mm, were separated from the remaining solidified material

Jan 1, 1963

By Douglas Valentim Fernandes Neves

O aço inoxidável austenítico 316 (316 SS) é um dos materiais metálicos mais utilizados em aplicações médicas devido à combinação de propriedades mecânicas e do baixo custo. No entanto, desgaste e corrosão estimulam mudanças permanentes no tecido humano que podem levar à perda do implante. O desenvolvimento de recobrimentos biocompatíveis e estáveis, que apresentem propriedades mecânicas adequadas, e que não sofram ruptura ou delaminação durante o uso, é uma solução muito promissora para melhorar o desempenho do 316 SS como biomaterial para implantes. Neste trabalho, uma liga intermetálica FeAlCr, que tem apresentado propriedades adequadas para a aplicação como biomaterial, foi utilizada como recobrimento sobre um aço inoxidável 316 de grau médico. O revestimento foi obtido utilizando-se a técnica magnetron sputtering por corrente contínua. Os parâmetros de deposição, como distância substrato - alvo e pressão de gás argônio foram mantidos fixos em 50 mm e 5 mTorr, respectivamente. As amostras foram preparadas com os seguintes valores de potência de sputtering: 50 W, 75W, 100 W e 150 W. A adesão dos filmes foi analisada por teste de scratch, enquanto a espessura e a taxa de deposição foram analisadas por perfilometria ótica. As características estruturais foram analisadas por microscopia eletrônica de varredura com EDS.

Aug 8, 2018

By David K. Hylbert

This study investigated stratigraphic and structural causes of roof falls in room-and-pillar drift coal mines in nine mines in eastern Kentucky. A contract modification provided for remote sensing studies in the Dunkard Basin of West Virginia and Pennsylvania. Results of stratigraphic and structural studies of an earlier contract (H0133018) were applied to several new mines. It was found that similar rock sequences contributed to unstable roof conditions. Structure contour and isopach maps were found useful In projecting these trends together with fence diagrams. Using Landsat imagery, a direct relationship was found between lineament direction and the trends of rock jointing, roof fall zones, "snap top" zones and sandstone channels in the examples cited. It was found that, by analyzing lineament intersections, from 50 percent to 65 percent coincided with roof falls.

Jan 1, 1980

By Mohamad Kharseh

There is general acceptance that climate change, which is the most important challenge facing humanity, is anthropogenic and attributed to fossil fuel consumption. Therefore, improving the performance of our existing energy systems and deploy more renewable energy resources is an urgent issue to be addressed. Geothermal refers to existing of heat energy in deep rock and sedimentary basins. This energy can be used to drive a power turbine to generate electricity. Traditionally, geothermal has been exploited in places with the plentiful hot water at relatively shallow depth. In the light of fact that ground temperature increases with the depth everywhere on the Earth, engineered geothermal systems (EGS) can be installed in any place to exploit the geothermal in generating energy. Unfortunately, the high exploration and drilling costs of boreholes is the main barrier to commerciality of EGS worldwide. In addition, there are technical problems associated with drilling big depth. In oil producing countries such problems can be addressed by utilizing whether active or abandoned oil or gas wells and, consequently, EGS can produce power at profit. The current study presents an analysis of a binary geothermal power generation system for commercial electricity generation in Qatar. For this purpose, two binary cycles are assumed the main difference between them is that the first one is air cooled while the other will be water cooled. The performance of the two cycles and the possibility of improvement has been shown. Economic analysis the power plant shows that the levelized costs of electricity is 3.6US¢/kWh and the pay-back time is less than 8 years.

Aug 1, 2013

By Werner Schwartz, Wolfgang Haase

Lead sulfide concentrates, which may include other lead concentrates, are sintered on an up-draught sintering machine without the addition of any diluting agents or fluxes. Subsequently they are melted in an oil- or gas-fired rotary furnace. The sintering and melting processes are based upon the following roast-reaction: PbS + 2 PbO = 3 Pb + SO, PbS + PbSO, =2 Pb + 2 SO, For obtaining a lead bullion free from sulfur, the sintering process is carried out in such a way that the sinter product contains a small amount of excess oxygen above that to react with the sulfides. At the end of the melting process, when the reactions are finished, the remaining small amount of oxide residues is reduced with coal to which a certain percentage of soda ash (about 1 pct of the lead bullion) is added. For the lead smelting process described neither coke nor fluxes—except soda ash—are required. This process is being utilized by a European smelter successfully and with a high lead recovery. The consumption figures for the smelting of 100 tons per day of lead concentrates are indicated. The lead content of the lead concentrates from modern ore dressing plants ranges from 65 pct to above 80 pct. In most lead smelters of the world these concentrates are smelted in a blast furnace. For blast-furnace smelting the concentrates have to be desulfurized and agglomerated by sintering. A requirement for the perfect operation of a down-draught sintering machine and of a blast furnace is a maximum lead content in the feed of 40 to 45 pct. For this reason, some lead concentrates have to be diluted by adding return slags, limestone, and possibly iron oxide and sand. As an example, 100 tons of lead concentrate with 72 pct Pb would contain 13.5 tons of gangue (including the zinc). To produce a perfect sinter with 42 pct Pb it would be necessary to add 70 tons of flux and return slag, more than five times the original weight of the gangue, to the sinter mix and blast-furnace charge. A correspondingly large amount of coke would be required in order that all of these materials reach the heat of formation and the melting temperatures of the slag (1200" to 1400°C) inside the blast furnace. The roast-reaction process presents a possibility for lead recovery without dilution of the concentrates. In this process the concentrate mixed with coal is placed upon a Newnam-hearth and air is blown through nozzles into the heated mix. AS a result metalllic lead and a relatively great amount of so-called .'Grey Slag" with a lead content of 25 to 35 pct are formed. The slag is sintered to eliminate sulfur and, after addition of the requisite fluxes, treatt:d in a blast furnace. Owing to the poor recovery of lead from the hearths and to the unavoidable heavy hand-work plus the risk of poisoning this process is utilized in very few 112ad smelters today. Since in mxny countries of the world coke is expensive and difficult to obtain, it appeared feasible to use the principle of the roast-reaction by modern sintering and melting methods with recovery of the lead in electric, or oil, gas, or coal-fired furnaces. Two processes are utilized on an industrial scale: A) Lead smelting in the electric furnace of the Bolidens Gruv A/B in Sweden, as described by S. J. Walldcn, N. E. Lindvall, K.G. Gorling, and S. Lundquist. B) The self-fluxing lead smelting of Lurgi Gesell-schaft fiir Chemie und Huttenwesen m.b. H., Frankfurt a M, Germany, which is described in this paper. In the Boliden process referred to above the sinter mix is pelletized by enveloping return fines with layers of flue dust, limestone powder, and dried galena concentrate. The roasting and agglomeration are carried out on a down-draught machine, and a slight excess of sulfur is left in the sinter product. During the smelting in the electric furnance the roast-reactions occur and a slag poor in lead and a sulfur bearing lead are formed. This latter is subsequently oxidized in a converter to obtain lead bullion and dross. The Lurgi-process achieves the maximum possible extent of the roasting reaction on the sintering machine. The wet flotation concentrates are blended with return fines (lead content 70 to 80 pet), any existing flue dusts and lead slimes—but without the

Jan 1, 1962

By F. N. Ward, E. H. Bailey

Camp and sample-site methods useful for determining about 0.5 to 16 ppm of mercury in soils and rocks have been devised to complement the analytical methods already widely used in geochemical prospecting. In the camp-site procedure the sample solution is obtained by digesting the finely powdered sample with hot nine-molar sulfuric acid and bromine, the latter generated in place. The solution is buffered at a pH of 4 and the mercury is extracted into a relatively small volume of an organic solvent. Estimations are made by comparing the color of the mercuric dithizonate extracted from the sample with that extracted from a standard. In the sample-site method the mercury is volatilized as the iodide and the latter is dissolved in a buffer solution from which the mercury dithizonate is extracted as in the camp-site method. The estimation is made by a similar comparison with standards or by comparing the color of the mercuric dithizonate with artificial standards prepared from Orange II, an azo dye. Except for the establishment of standards, the time required by an experienced analyst seldom exceeds 10 min for a camp-site determination and 5 min for a sample-site determination. Camp-site determinations are useful for commodity studies and intensive exploration programs; sample-site determinations are applicable to extensive and reconnaissance-type geochemical prospecting programs. The results obtained on samples taken along traverses above known mercury ore deposits in the California Coast Range and western Nevada demonstrate the usefulness of both methods in geochemical prospecting. The increasing use of the mercury content of soils, rocks, plants, or water in geochemical prospecting emphasizes the need for analytical methods which are sensitive and at the same time require a minimum of readily available equipment. Sensitive methods based on the absorption of ultraviolet light by mercury vapor are common in the literature31 12. Although none of these methods have been used to analyze the large number of samples required in geochemical prospecting, they could be extremely useful except for the expensive instrumentation usually required to make the methods quantitative Also, such methods are perhaps too sensitive for field use, especially around areas of possible contamination. Other methods involve the distillation of mercury from a relatively large sample and subsequent collection of the vapors in nitric acid6. The mercury is then estimated by titration with a suitable reagent such as thiocyanate. These methods are not trace methods because the actual amount of mercury measured is often as great as 50 to 100 milligrams. Also, such methods are unsuitable for geochemical prospecting because of the amount of equipment necessary and the time required. Still other methods involve colorimetry, and the sensitive reaction between dithizone and mercury (11).. is the basis for a number of reports on the determination of mercury in biological materials 5,7. Moreover, by the judicious use of solvent extraction and suitable complexing agents, the dithiaone method can be made almost specific for mercury. However, none of the dithizone methods have yet been applied to the determination of mercury in soils and rocks. In most colorimetric methods he sample solution

Jan 1, 1961

Computerized financial models can be used in all phases of the mineral industry from setting exploration targets to projecting profit plans for producing mines. In this presentation, I will use the example of an exploration company entering a new district to trace the application of financial models through the following intermediate phases. 1. Hypothetical analysis and strategy 2. First encounter of mineralization 3. Preliminary drill results - decision process 4. Feasibility studies 5. Post-feasibility negotiations 6. Producing mine 7. Acquisition of producing mines or mineral properties. I will conclude with a brief discussion of the mechanics of the financial models that we use at Wright Engineers Limited. As we progress through the seven phases, the availability and accuracy of the data will parallel the complexity and application of the financial models. However, the same basic financial models are utilized throughout the seven phases. I will begin by suggesting that the exploration group is somewhat autonomous and has the mandate to explore in Western Canada.

Jan 1, 1979

By D. J. Carney, E. C. Rudolphy

IT has been demonstrated that inclusion size, distribution, and composition affect the machin-ability of resulphurized steels. Merchant and Zlatinl concluded that large sulphide inclusions aided machining by forming a (lubricating) coating on the tool face. Boulger et al.² and Van Vlack³ noted that the size, distribution, and composition of the inclusions in the steel affected the machinability. Steel specimens containing large globular sulphide inclusions usually exhibited excellent cutting properties, while machinability was adversely affected by the presence of numbers of oxide-type inclusions. Consequently a thorough knowledge of all the factors which affect the inclusions in the final product is desirable. Since almost all the inclusions have their origin in liquid steel, it was necessary to begin a study of inclusions in free-machining steels by studying the inclusions and chemical segregation in the as-cast ingot. Very little information is available on the size, distribution, shape, and composition of inclusions in large, capped, free-machining steel ingots, particularly the B1113 grade. Gregory and Whiteley4 made a general study of the inclusions in a small, high sulphur, free-machining steel ingot. Also, numerous authors have described the solidification and segregation characteristics of the four basic types of steel ingots, namely, rimmed, capped,7 semikilled,7,8 and killed7,9,10 ingots. Most of these studies were made with plain carbon or low alloy, low sulphur steel. It was desirable to study not only the ingot but also the change in size, shape, and number of inclusions on rolling an ingot to a bloom and thence to a billet. This procedure was followed and it is hoped that this study may serve the dual purpose of adding to the general knowledge of ingot solidification as well as contributing to the knowledge of the size, shape, distribution, and composition of inclusions from the ingot to the billet in a high sulphur, free-machining steel. Procedure A 12.000-lb 23x35x75 in. slab ingot of the B1113 grade was cast, sectioned, and studied both macro-scopically and microscopically. An adjacent ingot from the same heat and of the same size was rolled to a 77/8x77/8 in. bloom and thence to a 21/2x21/2 in. billet. These various bloom and billet sections were also sectioned and studied macroscopically and microscopically. Sectioning the Ingot: The ingot herein described was obtained from the United States Steel Corp.'s South Works Bessemer Blow No. 0193, a B1113 mechanically capped heat. The 23x35 in. ingot (No. 2) was teemed according to normal procedures and after stripping and transportation to the rolling mill was not placed in the soaking pit but allowed to air cool in an upright position. When completely solidified, the ingot was cut into sections by means of a powder scarfing torch and further sectioned by saw cutting as indicated in Fig. 1. Cut No. 2 (1x10x12 in.) from sections A through H was cleaned thoroughly, macroetched in a solution of 50-50 water and hot muriatic acid and used to obtain a macrograph of a horizontal section from the surface to slightly beyond the center of the 23-in. ingot dimension. Cuts No. 5 and 3 (lx81/2xl0 in. each) from sections A through H were treated in a similar manner to obtain a macrograph of a horizontal section from the surface to slightly beyond the center of the 35-in. ingot dimension. The composite macrograph of these horizontal ingot sections, which shows a vertical section of the ingot from top to bottom, is shown in Fig. 2. It should be noted that sections No. 2 are normal to sections 5 and 3 in the composite. Drillings for chemical analyses were obtained from selected positions within the above-mentioned ingot sections as noted in Fig. 3. The oxygen content was determined by the vacuum-fusion method. Samples for microscopic examination were cut from

Jan 1, 1954

By S. Durapraj, R. T. Masethe, U. Maqina, A. C. Adoko

This study investigates seismicity and rockburst control within the Merensky Reef and UG2 orebody of an intermediate-depth platinum mine located in South Africa's Bushveld Complex. Utilising a mine-wide seismic monitoring system, over 1,900 seismic events were recorded between April 2023 and May 2024, revealing complex seismic patterns driven by mining-induced stress changes. The largest event, with a magnitude of ML 2.0, underscores the seismic risks associated with mining at increasing depths. Key findings include the spatial and temporal variability of seismic events, with high seismic activity observed near pillars and geological structures. A significant portion of events correlated with pillar and foundation failures. Moment tensor analysis identified dominant failure modes, such as stress-driven fracture closures and dynamic pillar failures. To mitigate risks, the study emphasises advanced support system redesign and rockburst risk assessment using the rockburst damage potential (RDP) method. The research highlights the importance of monitoring, geotechnical compliance, and adaptive designs in enhancing safety and reducing seismic hazards in underground mining operations. The study applied the RDP framework, classifying damage across R1 to R5 scales, where most instances fell into low-risk categories (R1 and R3), indicating minor to moderate damage. However, data limitations and the tabular stoping method's complexity led to inconclusive results, highlighting the need for more robust data and customised risk assessment tools. Despite these challenges, the RDP approach provided valuable insights into rockburst mechanisms and the effectiveness of existing support systems.

Jan 12, 2025

By F. H. Guise-Brown

Raise-boring experiences in the gold mines of the Anglo American Corporation Group by J. W. WILSON and P. C. GRAHAM Published in the Journal, October, 1972. Discussion by F. H. Guise-Brown., B.Sc. (Min. Eng.) (Member) INTRODUCTION Messrs J. W. Wilson and P. C. Graham are to be congratulated on an excellent paper containing several thought-provoking remarks. The concept of allocating a machine and equipment valued initially at some R300 000 on an experimental basis is courageous, and the results obtained could be of interest to the mining industry as a whole. The Gold Fields of S.A. Group approached this problem differently. In pioneering the start of this work in South African gold mines, they considered that expenditure of this magnitude would be warranted only if the proposition was viable right from the start. To this end, a Robbins R41 was purchased expressly to assist in the duplication of part of the No. 1 shaft system of the Doomfontein Gold Mine. Mr M. J. Koortsl has summed up the initial results of this work and given some preliminary costing, but inflation since that period has made these figures somewhat out of date.

Jan 10, 1972

By Richard Stoltz, Gary Smith

This paper presents the findings of a face ventilation investigation, conducted by the Ventilation Division, Pittsburgh Health Technology Center, at a underground coal mine operating in the Illinois No. 5 coal seam. The purpose of the investigation was to evaluate the effectiveness of a blowing face ventilation system in controlling methane liberation in the face area while mining with brattice distances up to a maximum of 40 feet from the face (extended cut). The continuous miner on this section was remotely controlled and equipped with a machine mounted dust collector. Data collected during this investigation included methane concentrations, intake and return air quantities in the face area, and scrubber air quantities. In addition, a time study was conducted to correlate the methane peaks to activities in the face area. In conjunction with the face study on Section A, separate methane liberation studies were conducted on three additional sections: B, C, and D. This was done in order to compare intake methane concentrations and methane liberations from other sections in the mine to the test section. Calculations indicated that the minimum intake air quantity necessary at the inby end of the line brattice for the section investigated was 7,000 cfm. Due to the higher intake methane concentrations and average methane face liberations measured on three other sections, the minimum air quantities necessary at the inby end of the line brattice were determined to be 13,900 cfm, 11,100 cfm, and 15,000 cfm, respectively.

Jan 1, 1991

By Joseph L. Patrick

Geologic exploration, which has experienced an upward trend in the United States since 1962, continued with increasing activity through 1965. New, spectacular discoveries of recent years and continued questionable economic and political climates in many areas of the world have continued to focus the exploration attention of an increasing number of companies on the United States. Traditional geologic field methods are being bolstered by development of new methods, techniques and geophysical instruments. Reports from the states emphasize that successful exploration is dependent upon the combined effort of geochemistry, geophysics and geology, with high quality laboratory support of the field effort becoming increasingly necessary. PROGRAMS NOW UNDERWAY IN WEST During 1965 mineral exploration reached an all- time high in Nevada. Although examinations are being carried on throughout the State, much of the work is limited to two areas. The first area to attract recent exploration interest, initiated by discovery of the Carlin gold mine by Newmont Mining Corp., is an area of windows near the leading edge of the Roberts Mountains thrust fault in northeastern Nevada. Included in this area are the old camps of Mountain City, Battle Mountain, Cortez, Mineral Hill and Eureka. Major mining companies are using a combination of surface mapping, geochemistry, geophysics (largely induced potential surveys), following up with drilling. The second area of exploration is along the Walker Lane, a wide shear zone that parallels the Nevada-California boundary from Virginia City to Las Vegas. Activity has been intense in the Tonopah-Yerington area. Discovery of a large copper-iron deposit following an airborne magnetometer survey first focused recent attention on Walker Lane. Sixteen major companies are involved in exploration in the State.

Jan 2, 1966

By W. T. Benson

The Bureau of Mines has examined selected deposits of titanium minerals throughout the United States. The titaniferous magnetite deposits in the western San Gabriel Mountains, Los Angeles County, Calif., were among those selected for study, and in 1956-58 field work was conducted to determine the extent and character of the titaniferous magnetite deposits occurring in the area. This report describes this work and summarizes earlier work in the area by the Bureau and others. The titaniferous magnetite occurs disseminated as seams and lenticular pods in an anorthosite belt 4 to 6 miles wide that trends about. N. 80° W., extending from sec. 13, T. 3 N., R. 12 W., to the west boundary of R. 14 W., at which point the titaniferous magnetite-bearing rocks disappear beneath the sediments in the Santa Clara River Valley.

Jan 1, 1962

By T. A. Oberhellmann

The Wobbler Feeder is a comparatively new materials handling and separating device. It is essentially a materials separating or scalping machine, particularly adaptable to separation of oversize particles from mine-run ore and to the separation of oversize materials from overburden, where the removal of oversize materials is necessary, as is the case in conveyor stripping. In construction, the Wobbler Feeder is simply a set of transverse bars of manganese steel, or other wear-resisting alloy steel, each having an elliptical cross-section with the major axes of adjacent bars displaced by ninety degrees so when rotating in unison, the pre-determined space between the bars remains constant.

Jan 1, 1959

By Leonid N. Shekhter, Corby G. Anderson, Esra Cankaya-Yalcin, Daniel G. Gribbin, Joseph D. Lessard, Larry F. McHugh

"A new process for smelting copper, Looping Sulfide Oxidation™ (LSO™) significantly and positively alters the smelting energy balance. This process uses copper oxide in conjunction with air for the desulfurization of copper and chalcopyrite concentrates. Copper oxide is regenerated in a flash type furnace and then ""looped"" back for the oxidation of the copper concentrate. Heat and material balances show that the Looping Sulfide Oxidation™ process requires significantly lower net energy than conventional technology. This process produces anode copper and higher S02 content off-gases for efficient formation of sulfuric acid. Additionally, the process yields slags with lower copper solubility as compared to the conventional technology.IntroductionBased on the positive results from LSO™ investigations [1], further lab investigations into co=on sulfides and metal oxides were performed. It was discovered that chalcopyrite (CuFeS2) is oxidized with copper (II) oxide. In this reaction over 98.5% of sulfur was removed as S02, while a significantly improved energy balance was observed. This paper intends to analyze both the conventional flash smelting process and the newly developed Looping Sulfide Oxidation™ process, comparing heat and material balance results. LSO™ yields improvements to the energy footprint for copper, opening the door to new copper processing opportunities."

Jan 1, 2013

1. Domestic Production and Use: In 1982, 13 companies operated 30 primary aluminum reduction plants, with 3 firms accounting for 55% of production. Washington, Oregon, and Montana accounted for 34% of the production; Texas, Louisiana, Missouri, and Arkansas 18%; Tennessee, Alabama, South Carolina, and North Carolina 14%; other central and eastern States 34%. Based on published producer prices, output of primary metal in 1982 was valued at over $5.5 billion. Aluminum consumption by 25,000 firms was centered in the East Central United States. Packaging accounted for an estimated 39% of domestic consumption in 1982, transportation 20%, building 14%, electrical 8%, consumer durables 8%, and other uses 11%.

Jan 1, 1983

By J. M. Karpinski, R. O. Tervo

A method and equipment have been developed for measuring the impact strength of grains of brittle materials. It is shown that brittle materials develop a characteristic particle size distribution when fractured by impact. A simple mathematical model has been found to describe this distribution, and one of the parameters of the model has been designated the r value. The r value is directly proportional to impact energy and, for fixed impact energy, it becomes a useful criterion of grain strength. Data are presented on the variation of the r value as a function of sample history, initial size, energy input and number of impacts. The results are interpreted in terms of the flaw theory of brittle fracture. The view is now widely held that brittle materials comprise a strong matrix, throughout which flaws or points of weakness are scattered at random. Failure always starts at one or more of these imperfections, and planes of fracture are established by the geometry of the flaw distribution. This view has been developed extensively by Weibull 1,2 and experiments of a confirmatory nature have been reported.3,4 Thus the process of fracture or size reduction of brittle materials can be regarded as a process of elimination or creation of flaws, and the concept of strength is intimately connected with the size and history of the specimen tested. One consequence of the point of view described above is that brittle materials will fracture to yield similar size distributions, although the impact energy to produce a given amount of fracture may vary greatly from one material to another. Generally speaking, experimental investigations of brittle fracture in solids have been confined to two extreme experimental situations. On one hand, we have tests on single specimens of well-defined, but artificial, geometry and, on the other, empirical data have been obtained on the feed-product relation in a variety of mills. The first situation necessitates specimen preparation, which in the case of brittle materials is equivalent to a priori size reduction, while the second one suffers from the disadvantage that little knowledge is gained about individual fracture events. Also, the statistical significance of single specimen tests may be questioned. Our approach lies somewhere between these two extremes. We test a large population of sized specimens and yet subject each specimen to a single fracture event. In 1956 G.H. Fetterley of Norton Co., Chippawa, Ont., made the empirical observation that product size distributions obtained from impacting sized grits can be described by the following equation where R is the proportion by weight of the grain that remains on a screen having an opening x on a side, x, is a parameter having the dimensions of length, and r is a number. This equation has since been derived on a theoretical basis by Gaudin and Meloy.5 They identify x, with the initial size of the test piece and r with the number of flaws per unit length. As will be shown below, our experiments indicate that both x, and r are functions of energy and hence they should more properly be called the effective initial size and the effective number of flaws respectively. Also, our empirical data give us no basis for assuming that r represents the effective number of flaws per unit length, rather than per unit volume. METHOD AND MATERIALS The impact test machine has been designed to feed specimens, essentially one at a time, into an evacuated chamber, where they fall freely and are struck with random orientation by one or the other of a pair of flat steel vanes. The vanes are mounted at opposite ends of a steel bar, which rotates at a closely controlled angular velocity about a vertical axis. Fig. 1 shows the flat circular vacuum chamber with the cover removed, and the horizontal steel bar with the steel vanes at the ends. The cylindrical pot in which the grains are collected after impact appears at the left. Fig. 2 shows the equipment assembled for a test. The vacuum chamber occupies the central position with the variable speed drive below the table surface, the vacuum pump at the extreme left

Jan 1, 1964