Search Documents

By John W. Walkiewicz

The U.S. Bureau of Mines has conducted studies to utilize rapid microwave heating to stress fracture ore samples. Iron ores containing hematite, magnetite, and goethite were subjected to microwave energy in batch operations at 3 kW and heated to average maximum temperatures between 840 and 940 °C. Standard Bond grindability tests showed that microwave heating reduced the work index of iron ores by 10 to 24 pct. In a microwave chamber designed to simulate a continuous throughput operation at 3 kW, the grindability of a taconite ore was improved by 13 pct at a bulk temperature of 197 °C. Because stress cracking occurred at a lower temperature, less energy was consumed. To further improve the economics of microwave fracturing, higher powers up to 16 kW were used to rapidly heat samples to relatively low temperatures in a continuous, belt-fed applicator. A significant improvement of grindability was obtained with a larger rod mill feed size in comparison to a minus 6-mesh Bond feed.

Jan 1, 2010

By J. B. Wagner, p. Hembree

The iron tracer diffusion coefficient of umstite has been measured at 110(fC across the phase field and at a single composition at 800°C. Assuming a simple cation vacancy model the tracer diffusion coefficient was found to be a linear function of the cation vacancy concentration at 1100°C. The equation is D = 3 x 20 29 where denotes the concentration of vacancies in numbers per cc. The tracer work at 800°C was carried out to investigate the reported "pinning" of tracer to the wustite surface at low temperatures. No evidence for the "pinning" of the tracer was found at 800°C in COz-CO gas mixtures. HIMMEL, Mehl, and Birchenall,&apos; Carter and Richardson,2 and Desmarescaux and La combe3 have measured the diffusion of iron tracer in wustite at several temperatures and compositions. The present work was undertaken to extend the measurements over a large composition range at 1100°C and to resolve certain apparent discrepancies in the data, expecially at lower temperatures. EXPERIMENTAL Wustite was prepared by oxidizing rectangular iron plates* in C02-CO mixtures. The samples were •The iron was supplied by the Battelle Memorial Institute courtesy of the American Iron and Steel Institute. The analysis is presented in Table I. quenched. Due to the inward flow of cation vacancies during oxidation, the center of the sample contained a thin void. The edges of the wustite slab were sanded until the sample could be split into two parts. Each part was then sanded on the front and back flat area until a smooth surface was obtained. The specimens were then replaced in the furnace and equilibrated at llOO°C in a predetermined COa-CO mixture by methods described elsewhere.4"6 The specimens were again quenched and the surfaces were lightly sanded to remove any roughness following the first equilibration. The specimens were then reequi lib rated in the same C02-CO mixture for thirty minutes in order to relieve any mechanical damage on the surface due to the polishing. The specimens were then quenched and the tracer was applied by an electroplating technique. The work of Carter and ~ichardson&apos; demonstrated that there was no systematic difference in the iron tracer diffusion coefficient in wustite if the tracer was plated, dried, or evaporated on the specimen. In the present study a piece of filter paper was saturated with an iron chloride solution of pH <* 3 that contained the tracer FeS5. The wustite was placed on the filter paper and made the cathode. A current density of 0.4 to 0.6 ma per sq cm was passed for about five to ten minutes. The thickness of the tracer layer was estimated to be about 7 x lom6 cm. This estimate was made by considering the area plated, the current flow, and time for plating and the activity of the iron in the plating solution. Different areas of the specimen were counted using a collimator to determine the uniformity of the tracer. Any specimen which exhibited a variation from the initial count rate (about 1500 cpm) by more than 15 pct was rejected. An estimate of the time necessary to convert the thin layer of iron tracer to wustite was made using the data of Pettit and wagner." he estimated time was 1 sec at 1100°C assuming linear oxidation kinetics. The shortest diffusion anneals were 1800 sec. The samples were suspended in the hot zone of a furnace by two platinum wires. Two separate specimens were run at the same time. Only the edges of each sample were in contact with the wires. The C02-CO gas of the same composition as that used in the pre-diffusion anneals flowed freely around the samples at a linear velocity of 0.9 cm per sec. To initiate a run, the specimens were lowered from the cold zone of a furnace to the hot zone by a magnetic lowering device." bout 60 sec were required for lowering. To terminate a run, the sample was withdrawn from the hot zone to the cold zone. Time zero for the beginning of the experiment was taken when the sample blended into the red glow of the furnace and conversely for the end of the experiment. The surface decrease method of measuring the tracer diffusion coefficient was used to collect the data. This method requires that counting geometry be reproducible because the specimen is counted before the diffusion anneal and after the anneal. A special jig was constructed for each specimen so the specimen could be removed from the jig and returned to the jig such that the well geometry was reproducible.

Jan 1, 1970

By E. A. Gulbransen, W. S. Wysong

The behavior of molybdenum and its surface oxides in oxidizing and reducing gas atmospheres and in high vacua at elevated temperatures is a question of scientific and technical importance. The use of molybdenum as a metal or as an important component in alloys in oxidizing atmospheres at elevated temperatures has been limited because of the unprotec-tive nature of its oxide film and the volatility of this oxide at moderate temperatures. This communication will present results of a vacuum microbalance studyl, 2 of the following problems: (I) the oxidation kinetics in the temperature range of 250°to 450°C; (2) the reduction with pure hydrogen of thin oxide films formed on molybdenum; (3) the volatility of the thin oxide films; and (4) the vacuum oxidation of molybdenum at high temperatures. Molybdenum is an interesting metal to study for the following reasons: (I) although the oxide-to-metal-volume ratio is greater than one, the metal is not considered protective; (2) the metal oxidizes readily at temperatures of 2 50°C and higher; (3) the oxides of the metal are reduced at temperatures as low as 500°C by pure hydrogen; (4) the oxide Moo3 melts at 795°C; and (5) the oxides of molybdenum are volatile at moderate temperatures. The following variables are studied in the oxidation kinetics: time, temperature, pressure, surface preparation, surface area, concentration of inert gas in the lattice, cycling procedures in temperature, vacuum effect and high temperature vacuum oxidation. The general interpretation of many of these factors has been previously discussed. Literature The oxidation of molybdenum has not been studied critically. Although many studies have probably been made on the kinetics of the oxidation process, none has been reported in the literature available to us. The crystal structures found in the oxidation of molybdenum have been studied by Hickman and Gulbransen." Their results showed that Moo2 is the oxide which is stable in contact with the metallic substrate throughout the temperature range 300 to 700°C. As the film thickens in the low temperature range, MoO3 becomes predominant on the surface. Above 400°C, Moos is no longer observed, MoO2 being the only oxide found. Hagg and Magneli6 have shown from an X ray study that the a-phase, orthor-hombic Moo3 is stable up to 1050°C. Two other phases, ß and ß with compositions close to MoO2.90 are also obtained. The formula for the 0 phase is not given but that of the ß phase is defined as Mo9O26 The 0&apos; phase is found to be built upon a monoclinic unit cell. A ?-phase is isolated from a preparation MoO2.75 heated for 17 hrs at 670°C. This phase is

Jan 1, 1949

By N. B. Wittman

The working of blast-furnaces is always of interest to many members of the Institute, for which reason I present an account of the working of the furnaces of the E. and G. Brooke Iron Company, of which I have had charge for the past four years. Although there are no phenomenal runs to record, there have been long periods of satisfactory work, and also periods—altogether too long—when one of our furnaces has behaved in a way which may be described as "nagging." All the conditions were apparently normal,but only a very light burden could be carried, which of course resulted in a high fuel-consumption to the ton of iron, as well as increased cost of labor. The two active furnaces of this company have the following dimensions and equipment : No 2. No 3. Feet. Feet. Height ....55 60 Diameter at bosh,.....- 15, 16 " hearth,...... 8 74 " stock-line,..10 91/2 Height of bosh,..23 23 " tuyeres,. 5 5 Diameter of bell,.7 6 Front,....closed open Both furnaces are blown with I. P. Morris condensing-engines, sim

Jan 1, 1890

By D. E. Walsh, J. T. Hollow, H. K. Lin, J. L. Oleson

At the Fort Knox Mine, Fairbanks, Alaska, carbon fines are generated during stripping, acid washing and kiln regeneration operations. Because of their gold value, carbon fines are collected by a filter press. Samples of the carbon fines show that they consist of 20% to 30% carbon, 70% to 80% grit and 0.17 to 0.68 g/kg (5.0-19.9 oz/st) gold. About 80% of the gold values in the carbon fines used in this study are in the form of recrystallized free gold grains, with the balance being the adsorbed gold on the carbon. High recoveries of gold and carbon were obtained from the fines in laboratory tests using froth flotation, with a high rejection of the grit. More than 95% of gold and 90% of carbon can be recovered by three stages of gold flotation followed by two stages of carbon flotation. Key words: Gold recovery, Activated carbon, Flotation

Jan 1, 2003

By H. Wagner

INTRODUCTION One of the most important tasks facing the opera¬tors of deep-level gold mines is supporting the roof strata in the immediate vicinity of the stope faces. Some of the deeper gold mines operate at depths ex¬ceeding 3.0 km; the extreme mining depths and the tabular nature of the gold-bearing reefs result in high stress concentrations at the abutments of the stope faces. These, in turn, lead to extensive fracturing of even the hardest quartzites. Although most of these fractures develop in a stable manner, there have been many instances where rock failure occurred suddenly, resulting in a considerable amount of damage to the underground workings. Such violent rock failures commonly are known as "rock bursts," and they are typical of many hard-rock mines that are operating at great depths. To safeguard mining personnel and equipment against the effects of stable and unstable rock failures, adequate support must be provided to the fractured roof as quickly and as effec¬tively as possible. Traditionally, timber props and timber packs have been utilized to provide roof support in gold mine stopes up to 1.5 m wide. Both methods of support suffer from the fundamental disadvantage of generating sup¬port only in reaction to movement of the strata. Since unrestrained strata movement often results in a break¬down of the frictional forces that preserve the stability of the intensely fractured roof strata, it is essential that stope supports have rapid load-bearing characteristics. An equally important requirement is that the roof sup¬ports be able to accommodate displacements of the rock mass. A third requirement of roof supports in gold mine stopes is that they must be able to accommodate the high rates of stope convergence typical of rock bursts. The hydraulic prop is the obvious choice for sup¬porting the stope faces in deep-level hard-rock mines. Since hydraulic props can be installed with a positive setting load, they can provide roof support that is independent of strata movement. By equipping the hydraulic props with yield valves, they meet the require¬ment of being able to accommodate rock-mass displace¬ments. However, the requirement of being able to accommodate high rates of stope convergence without overloading or damaging the rock strata was rather difficult to meet. The first introduction of hydraulic roof supports into gold mine stopes showed that conventional hy¬draulic props, such as those used in coal mines, were completely unsuitable for a number of reasons, in¬cluding: 1) Since drilling and blasting are used to break the ore in gold mine stopes, supports installed close enough to be effective are exposed to the vigor of the blast. Conventional coal-mining props suffered extensive blast damage, despite being protected by a blast barricade. 2) The environmental conditions in gold mine stopes favor corrosion. The high quartz content and the abrasiveness of the ore further aggravate the problem by damaging the protective coatings on the hydraulic surfaces. Simply changing the design of gold mine props to have the piston project from the bottom rather than from the top resulted in dramatic improvements in the service life of the props. The improvement was attributable to external water in the stope washing the fine quartz grit away from the hydraulic surfaces, rather than concentrating the particles between the cylinder and piston of the prop. 3) The most obvious deficiency of conventional coal mining props was their inability to yield rapidly enough to accommodate the high convergence rates of rock bursts without overloading the supports. As a result of the first disappointing results with hydraulic supports in narrow hard-rock stopes mined by drilling and blasting, a comprehensive research and development program was undertaken. The hydraulic props resulting from several years of research and exten¬sive field trials are described herein. To date (1978), more than 130,000 of these props have been manufac¬tured and used successfully in a number of mines and under a variety of geological conditions. Although many of the detailed specifications are directed spe¬cifically to gold mining conditions, there are a number of points that are of general importance in the design of underground support systems for tabular excavations in hard-rock mines of any type. SPECIFICATIONS FOR RAPID-YIELDING PROPS Yield Requirements The yield requirements for hydraulic props in deep¬level hard-rock gold mines are summarized in the fol¬lowing paragraphs. Slow Yield Load: The design of the yield valve must allow the yield load to be preset to between 300 and 400 kN. This flexibility is necessary to account for the differences in the strength properties of the roof and floor strata. Drop in Load Under Slow Yield: The drop in sup¬port load during a slow yield should be as small as possible, but it should not exceed 5% of the slow-yield load. Rapid-Yield Load: The load exerted by the prop during periods of rapid stope convergence should be as close as possible to the slow-yield load. The rapid-yield load should not exceed the slow-yield load by more than 15% at yield rates of 1.0 m/s. The limitation of the upper valve of the yield load is important; it deter¬mines the hydraulic performance of the yield valve. The hydraulic props that satisfy this criterion have performed well, even under severe rock bursts. Load Changes Under Rapid Load: On some of the earlier rapid-yielding props, conditions of dynamic in¬stability were observed; these conditions resulted in a complete breakdown of support resistance. It is essen¬tial that the force oscillations during rapid yielding be as small as possible, not exceeding -*5% of the average rapid-yield load at a yield rate of 1.0 m/s. Longitudinal Resilience: During rock bursts, vibra¬tions of the roof and floor sometimes are observed. To

Jan 1, 1982

By XUEWEN WANG, Xingming Wang, Xuehui Liu, Mingyu Wang

The Panzhihua titanium concentrate contains high contents impurity of Mg and Al, and it is difficult to prepare high grade titanium slag. To solve this problem, a sodium salt roasting-acid leaching process was proposed for removing the impurities in titanium concentrate. Na2CO3 and titanium concentrate were mixed with the proportion of 1:4, and then the mixture was roasted at 700 for 2h. After roasting, the calcine was leached with HCl solution. The experimental results showed that the removal rate of silicon and aluminum can reach more than 85% and 80%, respectively, at leaching temperature of 70 , leaching time of 4h, HCl addition of 5ml and liquid-solid ratio of 4:1.

Jan 1, 2016

By T. T. Chen

The iron- and elemental sulphur-rich residues generated during batch and continuous pressure oxidative leaching of Inco Limited&apos;s Voisey&apos;s Bay nickel sulphide concentrate at 150°C in sulphate-chloride media have been characterized by X-ray diffraction analysis and scanning electron microscopy. The residues produced during the continuous leaching of the pentlandite- and pyrrhotite-rich concentrate consisted predominantly of hematite and elemental sulphur, with minor amounts of goethite. Although the iron species generated in the batch leaching tests were similar to those formed during continuous leaching, the morphologies of the hematite and goethite particles were significantly different. In addition to the tiny spheroids and bubble-like particles, the hematite and goethite generated during batch leaching occurred as larger "hollow" shell-like particles that contained residual pentlandite or sulphur cores. Although the leach solution from which the iron oxides had precipitated contained 5-15 g/L sodium, sodium jarosite was detected only in trace amounts in a few of the leach residues.

Jan 1, 2006

By Y. H. Liu, R. Kiessling

The chromium, iron, and tungsten borides have been treated with ammonia at different temperatures. They are attacked, forming metal nitride and boron nitride, and the results are summarized in the tables. In the tungsten-nitrogen system a new phase has been observed, closely related to the known ß phase. THE present investigations were begun to determine whether any ternary phases exist in the transition metal-boron-nitrogen systems for the metals chromium, iron, and tungsten. Attempts to prepare such phases were made by using the borides of the metals as starting materials and ammonia as the nitriding agent. Although no new ternary phases have been found to exist, the results may be of some interest with regard to the technical use of the borides. General Procedure A steady stream of dry ammonia was passed through a silica reaction tube, the central part of which was kept at the reaction temperature while its ends were water-cooled. Weighed amounts of the borides were placed in a silica or porcelain boat and heated at different temperatures for a period of 10 to 24 hr. The boat was then rapidly removed from the reaction zone to the quenching zone, the water-cooled end of the tube, by means of a simple magnetic device. The nitrogen content of the products was determined by the increase in weight of the specimens, and the phase analysis was carried out by X-ray methods. For this analysis a Guinier-type camera, constructed at this Institute, was used with a bent, ground quartz crystal monochromator. The general reaction, which occurred in all the experiments, may be summarized as: metal boride + nitrogen -* boron nitride + metal nitride For the higher temperatures pure metal was formed instead of metal nitride. It is sometimes difficult to Discussi detect boron nitride among the reaction products by X-ray methods. Prolonged exposures of some specimens showed, however, that the strongest reflections of boron nitride were present. Additional evidence for its presence was given by the increase in weight of the specimens after nitriding. This increase was always greater than that corresponding to nitride formation only and was often in close agreement with the formula given above. Finally, a white powder, which was shown by X-ray methods to be boron nitride, was left if the reaction products after nitriding Fe,B at 448°C and FeB at 768°C were dissolved in diluted sulphuric acid. Ch rom iu m - Boron - N itrogen Five phases have been found to exist in the Cr-B system,&apos; , and all were used as starting materials. Investigations on the Cr-N system have shown the existence of two nitrides with ideal composition Cr,N and CrN.".&apos; * The results are summarized in Table I. All the borides are attacked and decomposed into chromium nitride and boron nitride, and the kind of nitride formed depends only on the temperature and not on the boride used as starting material. The lowest reaction temperature, however, is different for the different borides. Thus the 6 phase begins to react at about 735 "C, the r phase at about 800°, the S phase at about 900°, the 7, phase

Jan 1, 1952

By Kenichiro Yamamoto, Masaharu Murakami, Toshio Matsuzawa, Yoshiyuki Ikeda

In a series of tests, we examined the dynamic shock resistance of Emulsion Explosive in water and compared it with that of Dynamite and Slurry Explosives. Both the distance between donor and acceptor charges and the delay initiation timing of the acceptor charges were varied. The resistance to dynamic shock varied with the properties of the explosive. In the case of emulsion explosives, there seems to be a certain proportional relationship between the critical detonation distance and the delay initiation timing. Some blasting results with emulsion explosives are described.

Jan 1, 1983

By R M. Lonsdale, M D. Goodz

"With an ever-increasing dynamic business environment, successful stakeholder engagement is becoming more fundamental to achieving desired outcomes in today’s mineral resource industries. It is imperative to be proactive in correctly identifying all stakeholders and initiating early contact. This builds trust relationships, which are essential to ensure future decision-making processes are streamlined.Often neglected is a company’s obligation to select the correct team to represent its values and mission. The company requires representation that demonstrates an understanding and commitment to its operations and people, the safety and health of the community and environment, and most importantly, represents the decision makers. Stakeholders need to believe that the engagement process is genuine and that the company representatives are aligned towards targeted outcomes that will benefit all stakeholders in some way.Among others, stakeholders include shareholders, staff, families, community, service and supply providers, regulators, land managers and the political/financial framework. Depending upon the mineral product, stakeholders may include ports, international offtake clients and those along the route. New stakeholders continually evolve during mine life, and the engagement process requires an investor–community–public relations (IR/CR/PR) team to anticipate change and actively scan perception and chatter.Successful communication is more than delivering a message; it is about engaging stakeholders and confirming the feedback loop to ensure the correct message is communicated and the relationship is strengthened.The nature of the mining company’s branding and the public’s perception of company intentions will dictate whether there is willingness to allow the company to follow its desired path. This in turn controls the company’s most scarce resource – time. Timing is fundamental to achieving targeted outcomes with minimal cost and delays, marketing in a peak cycle and acquiring critical capital for development and growth.Branding and stakeholder engagement effectively controls and extends to the company its social licence to operate.CITATION:Goodz, M D and Lonsdale, R M, 2016. Stakeholder engagement – who are you, what do you stand for and why? In Proceedings International Mine Management Conference, pp 239-248 (The Australasian Institute of Mining and Metallurgy: Melbourne)."

Jun 22, 2016

By F. Vallianatos

The main objective of the present investigation is to demonstrate the potential of ERT (Electrical Resistivity Tomography) and HVSR (Horizontal to Vertical Spectral Ratio) methods as geophysical tools for the monitoring of subsurface cavities. The survey sites are located in Akrotirion Peninsula, almost 7 km east of Hania town in Crete Island, Greece. At this site, the surface expression of a known cavity with a diameter of about 25 m is visible. The true geometrical characteristics of the cavity were unknown since there are no boreholes in the area. The focus of this survey was primarily of a qualitative nature and the objective was to determine if ERT and HVSR could be used to identify and map near surface cavities (less than 30 meter in depth) in a limestones kartsic area such as Akrotirion Peninsula. A 2D/3D model of the resistivity distribution is given for the area under investigation. Furthermore, analysis of microtremors data provides a profile of the seismic response of the tectono-karstic voids.

Jan 1, 2006

By Ronald S. Conti

A new U.S. Bureau of Mines 6.8-L ignitability furnace was used to study the thermal autoignitability of dust clouds. This furnace has a quartz window to allow observation of the early stages of the ignition process and to allow measurement of the dust explosion temperature. Thermal ignitability data were obtained for various carbonaceous and metal dust clouds, with particular emphasis on various ranks of coal dusts. One of the reasons for the construction of the new 6.8-L furnace was to evaluate the effect of chamber volume on the measured autoignition temperatures. Therefore, data from the 6.8-L furnace were compared with data measured earlier in a 1.2-L furnace. The conclusion is that the autoignition temperatures measured in the 6.8-L furnace were only slightly lower on average than those from the 1.2-L furnace. These data on the minimum autoignition temperatures of various materials will be useful in analyzing the thermal ignition hazards of dust clouds in the mining industry and other industries that manufacture, process, or use combustible dusts.

Jan 1, 2010

By C Zschiesche, alam, A Abadias Llamas, A K

An integrated experimental and thermodynamic modelling study of the phase equilibria in the FeOFeO1.5- SbO1.5-SiO2 system in equilibrium with liquid Sb and Sb-Ag metal has been undertaken to develop thermodynamic model of the Na-Si-Fe-O slag system with S-Sn-Sb-Pb-As minor elements based on the experimental investigation of the selected sub-systems. New experimental phase equilibria data at 575–1600°C were obtained for this system using high-temperature equilibration of synthetic mixtures with predetermined compositions in sealed silica ampoules, Fe3O4 basket in sealed silica ampoule, or Re foils, a rapid quenching technique, and electron probe X-ray microanalysis of the equilibrated phase compositions. Experimental results on phase equilibria in the Fe-Sb-Si-O+Sb metal system obtained helped to produce initial set of model parameters in the slag. The primary phase fields of quartz/tridymite/cristobalite (SiO2), 2 high-SiO2 immiscible liquids, olivine (Fe2SiO4), wustite (FeO1+x), spinel (Fe3O4), schafarzikite (FeSb2O4) and valentinite/ senarmontite (Sb2O3) were identified in the binary Sb2O3-FeO, Sb2O3-SiO2 and ternary FeO-Sb2O3- SiO2 systems. Further refinement, particularly the distribution of excess Gibbs energy among the ‘FeO’-Sb2O3-SiO2 and ‘FeO1.5’-Sb2O3-SiO2 interactions, required the study at more oxidising conditions. Oxygen potential was increased using an inert metal (Ag) rather than using the gas flow due to the high volatility of Sb-containing gaseous species. Liquid Ag-Sb alloys with low concentrations of Sb of a few percent were used. At these conditions, the concentration of AgO0.5 in slag was below 1 mol per cent, which had a small effect on observed phase equilibria in the system. In addition, important information on the solubility of silver in slags is produced, which can be valuable when treating electronic recycling materials or primary silver-containing ores through the Pb process. Two main areas were addressed: the spinel-tridymite boundary (40–70 per cent SiO2 in slag) and spinel liquidus near 15 per cent SiO2 in slag. This allowed to deconvolute the separate contributions of the FeO-Sb2O3(-SiO2) and Fe2O3-Sb2O3(-SiO2) systems.

Jun 17, 2024

The compilation of the sampling nomogram is an essential tool in the optimization of a sampling protocol and allows the operator to track the change in relative variance of the sampling error (the Fundamental Sampling Error) during steps of comminution and mass reduction in the process of recovery of the aliquot for analysis. The data required for the compilation of the nomogram are centered on the sampling constant K and the exponent, alpha (a), of the nominal fragment size. Five methods for establishing K have been identified, but two of them, the duplicate sampling analysis (DSA) method and the heterogeneity test (HT) method purport to provide such a constant. The step by step procedure for each of these two methods was examined using broken ores from three mines, Mponeng mine, Kloof mine and Lily mine. Two of these mines (Mponengand Kloof) are typical Witwatersrand-type gold mines, while the third is an Archaean-type, shear related gold deposit situated in the Barberton Greenstone belt. The results suggest that the factor K derived using the heterogeneity test is very accurate, but applies only to a single size fraction in the spectrum of comminuted materials. The calculation of gold grain sizes using this method appears to be too small and does not conform with the results from detailed mineralogical studies. The factors K and alpha derived using the duplicate series analysis, are appropriate for use across a wide spectrum of comminution sizes and also provide realistic gold grain sizes, comparable with the equivalent circular diameter of gold grains identified in mineralogical studies of the ores.

Jan 1, 2009

By Niall Weatherstone

Introduction Risks & Rewards of a Mineral Reserve Life History of a Mineral Reserve Estimate Risk Assessment & Management Mineral Resource & Reserve Estimation Good Practice Mineral Resource & Reserve Reporting Conclusions

Jan 1, 2003

By F. J. Tuck

THE Hayden mill of the Ray Mines Division, Kennecott Copper Corporation, is now equipped with two 9 by 12-ft. rod mills having a daily capacity of 6000 tons of coarse-crushing plant product from the mine. These two mills crush a ¾ -in. product to approximately 19 per cent on 14 mesh, with only one pass through the mills. Each rod mill has taken the place of four sets of 42-in. Garfield rolls with their attendant screens, and is making a product for the fine-grinding department comparable to that formerly made by rolls in closed circuit with screens. The rod mills are lined with wave-type liners of white iron, varying in thickness from 2 to 3 in. The shell and end liners last approximately 600 to 700 days. Manganese-steel lifter bars, 3 5/8 in. thick, are used, spaced about 11 in. apart, and they give a life of 350 to 450 days. Fig. 1 shows the general design and thickness of the wave-type liners and lifter bars; 1 ½ -in. liner bolts are used to fasten the lifter bars to the shell. The shell liners are held in place by the tapered lifter bars. A 3/4-in. wood backing is used between the shell and the liners. The rod-mill scoop has removable manganese-steel lips, which last 250 to 300 days. The scoop proper has shown some wear in spots, and these spots have been protected by rubber bolted over them. Water and milk of lime are added to the elevator feeding the mill, but not enough water to cause a pounding of the rods against the shell. Roughly, the water added amounts to 25 to 33 per cent of the ore. Additional water is added to the mill discharge to facilitate distribution to the classifier bowls in the fine-grinding department. Rods are charged to the mill once every four days, the mill being down about 20 min. for this purpose. Other delays are negligible, and usually are due to a loose liner bolt or possibly power trouble. However, these occasions are extremely rare. In charging the mill, a crane is used

Jan 1, 1938

The present paper records a chapter in the history of the development of an electrolytic manganese industry in the United States.l A relatively large pilot plant at Boulder City, Nev., for the production of electrolytic manganese was conceived as part of the national defense program to make this country independent of foreign importation of manganese ore. The first authorization, in November 1940, however, was for a plant of only several hundred pounds capacity. Improvements in process and subsequent additions to some sections authorized by subsequent appropriations have enabled the plant to produce as much as 2000 lb. of manganese a day for substantial periods of time. Since in pilot-plant operation frequent changes are made in the flowsheet, the electrolytic manganese plant was constructed so that equipment could be rearranged with a minimum of expense and time. Several different kinds of the various ¦ types of equipment were installed for many of the operations, so that the one most suitable for the process might be chosen. Pilot-plant operations were begun in November 1941 and have been continuous except for brief shutdowns for installation of new equipment. More than one-half million pounds of manganese have been produced. The improvements in process, selection of equipment, and operational procedure developed are recorded here. Development of the Process The work of early investigators on the electrolysis of manganese solutions has been reviewed by Shelton. Until the method of continuous production of manganese by electrolysis was perfected by SheltonZm4 and coworkers, no practical procedure had been devised. The method as originally developed by the Bureau of Mines will be described here only briefly. The manganese dioxide ore was reduced in a rotating kiln-type furnace at 650°C. in an atmosphere of "city" gas. The calcine was leached with spent electrolyte and enough sulphuric acid to produce a neutral leach solutioil containing 2 5 grams manganese per liter as manganese sulphate and zoo grams per liter of ammonium sulphate. Arsenic was removed from the solution by the addition of 0.0j gram iron per liter as ferrous sulphate followed by oxidation with manganese dioxide, and aeration, which precipitated the iron as ferric hydroxide. The solution, free from iron

Jan 1, 1944

By F. A. Schaufelberger

METAL can be recovered from a leach solution either indirectly by precipitation as a compound that is later reduced or directly by electrolysis, cementation, or chemical reduction, for example, with hydrogen. The widely used electrowinning processes suffer from the inefficiency of converting fuel to low voltage dc and from low current efficiency when complete recovery is attempted. With the batch hydrogen reduction methods outlined here, a unit of fuel converted to hydrogen in modern gas reform plants will make two to three times as much metal as can be obtained with electrolysis. With continuous hydrogen reduction, four to six times as much metal is obtained. Precipitation by reducing a metal salt solution with H, has been known for almost 100 years.&apos; Commercial use of the process, however, awaited the research and development program initiated by Chemical Construction Corp. Within the last few years this program has brought about construction of commercial plants, listed on page 545. Ideal hydrogen reduction will precipitate a pure metal from a solution obtained by commercial leaching methods at a rapid rate without excessive temperatures and pressures. The metal precipitate will be of desired size and density, less than a percent left in the vessel being deposited on the wetted parts. The present discussion will outline Chemical Construction Corp.&apos;s early development program and will discuss the chemistry and mechanics of reducing copper, nickel, cobalt, and cadmium from solution by H2. Work on selective reduction of nickel from cobalt has been described earlier.&apos; The article is not concerned with precipitation of copper by disproportionation of cuprous solutions where yield is limited3 to 50 pet.* It does not discuss use of gaseous reducing agents other than hydrogen, such as SO2,4 which may lead to contamination with sulfur, or C0,5 which is considerably more expensive than hydrogen and produces a gaseous reaction product, CO2.† The article does not include use of other nongaseous reducing agents,7 which are far more expensive than CO. Also, notes on the history, engineering design, and performance of commercial plants have been given elsewhere.&apos; It will be shown that a pure metal can be precipitated by reduction with hydrogen from solutions obtained by commercial leaching methods when the solution composition is controlled and proper acidity, proper metal ion concentration through controlled complex formation and hydrolysis, and adequate agitation for suspension of metal and for transfer of reducing gas are maintained. Reasonable temperatures and pressures are used which are less excessive than those used by previous workers, although they are maintained at above equilibrium values because of process kinetics.&apos; It has also been found necessary to induce nuclea-tion and to control growth and agglomeration in order to get a powder product of suitable physical properties. Review of the Literature Muller, Schlecht, and Schubardt of I. G. Farben10 claimed a successive reduction of silver, copper, nickel, cobalt, and zinc from ammoniacal solution by applying progressively higher temperatures and hydrogen pressures. Metals produced by this technique were not pure, since no attempt was made to adjust the solution composition between the different reductions. It is doubtful that the zinc product contained any metallic zinc. The major contribution to the literature" was made by the Ipatiews,12 who prepared platinum, iridium, copper, nickel, cobalt, lead, tin, arsenic, antimony, and bismuth at often extreme conditions, up to the critical temperature of water (373ºC), at 1500 to 7500 psi, for as long as several days. Even cadmium and zinc were claimed to have been observed in trace quantities precipitated together with basic salts. In the early experiments solutions were not stirred. Extensive formation of oxides and basic salts occurred, probably long before sufficient hydrogen could be supplied in the unstirred liquid to decrease the metal concentration in solution by reduction. The importance of hydrogen pressure was

Jan 1, 1957

By E Marambio, R Brändle, G Fischer, G von Rickenbach, M Fuentealba, L Burgos, R Luis Fonseca

In underground excavations, under high stress conditions, ground support systems must provide safe and effective response to dynamic events. These systems must be capable to resist dynamic impacts and yielding during the loading process. In this context dynamic testing of the reinforcement and retaining elements that compose the ground support system are required to study and improve the behaviour of these elements under dynamic load events. During the last years, Geobrugg has been working on the improvement of retaining products by testing them in a large-scale impact test facility located at Walenstadt, Switzerland. The test facility is composed of a double level platform with a square-shaped pyramidal trunk geometry, in the upper level housing a loading mass that drop from a height up to 5m. The loading mass is guided by one central steel pipe, and the impact occurs in the sample to be tested located at the lower level in a slab with an area of 3.6m × 3.6m where the ground support system is installed. During the last years, this innovative facility has been used to test several configurations of ground support systems. The results of these tests have enabled to improve the understating of the behaviour of ground support systems under dynamic loads. In this manuscript, the back analysis through numerical methods (central difference and Newmark’s method) of a ground support system tested with a double dynamic impact in 2019 is presented.

Apr 26, 2022