Search Documents

By G. S. Dobby

The Cu-Au flotation cleaning circuit at Freeport Indonesia consists of three stages: two flotation column stages, in series, followed by one stage of mechanical cells. Concentrate from both column stages is combined as final concentrate. The mechanical cell stage has two parallel rows of seven Wemco 500's. The Wemco concentrate is recycled to feed of the second column stage and Wemco tails report to final tailings. This paper examines cleaner circuit performance, focusing on size-by-size performance of the individual stages and the individual mineral groups -copper sulfides, gold and in sol minerals. The recovery mechanisms of insol minerals are examined.

Jan 1, 1995

By Stephen H. Stow

Economically important heavy minerals have been studied in the sediments of offshore Alabama and Mississippi. Sandy samples have been collected from a 200 square mile area from Dauphin to Horn Islands. The minerals concentrate in sediments under 10-35 feet of water and may reach 2.4% by weight. The distribution of total heavies is related to the hydrodynamics of offshore Alabama-Mississippi; the greatest concentrations are in the Gulf where the longshore current is diverted from its westerly flow. Preliminary observations indicate that the distribution of individual minerals can be correlated with wave energy, particle size, specific gravity, and morphology of the mineral grains. The coarser, lighter minerals--kyanite, sillimanite and staurolite--are concentrated in relatively low energy areas in the Mississippi Sound and at the end of Dauphin Island in the Gulf; maximum concentration for each is 34.5%, 6.0% and 43.2%, respectively, of the heavy minerals. The smaller, more denseminerals--ilmenite, rutile, leucoxene, zircon and monazite--are concentrated in high energy areas in the Gulf; maximum concentration for each, respectively, is 42.0%, 5.9%, 9.652, 10.7% and 1.3% of the heavy minerals. On the basis of a minimum concentration of 1% heavy minerals, two areas have been identified and the total amounts of different minerals have been calculated to a depth of 25 feet of sediment. The data are preliminary and in no way indicate recoverable reserves, but they do show that on the order of 20 million tons of heavies exist. These deposits should be termed "indicated resources".

Jan 1, 1975

By W. J. Schlitt

Introduction Heap and dump leaching are low-cost metal production techniques that are gaining in popularity among gold and copper producers. However, the flow of solution in heaps and dumps has received little attention in the literature. This is unfortunate since solution flow is one of the few parameters subject to operator control. Thus, solution management may well influence both operating costs and plant performance. Costs aside, there are two important aspects of solution flow to be considered -the metallurgical and the hydrological. Some of the metallurgical factors have recently been discussed (Jackson and Ream, 1980; Schlitt, 1984; Schlitt and Nicolai, 1987). These cover solution application rates and methods, irrigation rates, leach-rest cycles, and nonvertical solution flow. According to Caldwell and Moss (1985), however, the hydrological aspects may be more significant than the metallurgical ones. In particular, a phreatic surface will form when excessive flow leads to an accumulation of water and the associated buildup of pore pressure within the rock pile. Such internal flooding can either originate at the foundation or at some other zone of very low permeability. The latter condition gives rise to an impounded volume of solution, i.e., a perched water table (see Whiting, 1985). Caldwell and Moss point out that a rising zone of saturation is probably the most common cause of dump failures. Of course, such failures have even occurred in heaps that were carefully prepared for leaching (Milligan and Engelhardt, 1984). Thus, flooded conditions and perched water tables represent an important safety consideration as well as having an impact on metallurgical performance. The following sections describe a case history in which a perched water table developed within a copper leach dump. The description includes background information, solution flow rates, and metallurgical data. Then this situation is compared to one involving normal drainage. Description of the leach system The leach dump in the case history is located at a shutdown open pit copper mine. It was built in two lifts, with the second added some 20 years after the first. There is little detailed information available on the initial lift. It was built with rail¬hauled waste and was generally less than 30 m (100 ft) high at the crest. The area available for leaching was about 120 m (400 ft) wide and more than 380 m (1250 ft) long. The dump surface was prepared for leaching by dozing ponds approximately 12 m by 12 m by 2.5 to 3.0 m deep (40 ft by 40 ft by 8 to 10 ft). The ponds were leached by flooding with barren leach solution returned from a scrap iron cementation plant. Based on mill feed at the time, the average waste grade was probably close to 0.4% Cu. The first lift leached well. It accepted high flows, which together with the waste grade, produced a rich pregnant leach solution (PLS). Old records indicate a PLS of "50 lb Cu/ 1000 gal," or about 6 g/L. As later drilling would indicate, such a high tenor led to dissolution of considerable scrap iron that was returned to the dump. The iron then hydrolyzed and settled out on the pond bottoms. In addition, the waste settled substantially so that the unleached crest was 3.0 to 4.5 m (10 to 15 ft) above the ponded area. Eventually, the leachable copper was extracted and the dump became less permeable. Thus, the PLS tenor dropped until it became uneconomical. About ten years later, mining resumed and the decision was made to add another lift to the dump. This was done without giving much thought to a subsequent leach operation. Hence a 24-m (80-ft) lift was built on top of the original dump. The surface of the latter was not prepared in any way, e.g., by leveling and/or deep ripping, prior to over-dumping. Examination of subsequent drill cuttings indicated that the new lift contained about 0.2% Cu, with chalcopyrite and chalcocite (50:50) being the predominant copper minerals. Most of the chalcocite occurred as rimming on the abundant pyrite, with the pyrite to chalcopyrite ratio estimated at 10 to 1. The use of 100- and 120-ton trucks for haulage caused some waste compaction during emplacement. In addition, the host rock itself was relatively soft, being a porphyry intrusive material that was partially altered to clay. As a result of the initial compaction and clay swelling, the rate of water percolation from the new leach ponds was slow and the ponds often contained considerable standing water. Even frequent ripping failed to provide a sustained improvement in the percolation rate. The poor surface permeability was exacerbated by hydrolysis of iron salts which settled as a layer on the pond bottoms. Partly as a result of the permeability problems, metallurgical performance was not up to expectations. These had been based on laboratory tests which showed about 20% copper solubilization in two weeks. Continued copper extraction in the tests also suggested that a substantial percentage of the copper would eventually be recovered. However, the PLS grade in the actual operation peaked briefly at about 0.48 g/L Cu (4 lb Cu/1000 gal), then declined to a range of only 0.24 to 0.36 g/L Cu (2 to 3 lb Cu/1000 gal). The poor leaching was traced to a lack of oxidation of the sulfides. There were two principal observations supporting this conclusion. First, there was no evidence of any heat being generated within the dump. As discussed elsewhere (Schlitt and Jackson, 1981; Hiskey and Schlitt, 1982), pyrite oxidation is quite exothermic and the high pyrite content of the waste should have led to an increase in the temperature of the leach solution as it percolated through the rock pile. Second, there was no sign of any natural convective air flow through the dump.

Jan 1, 1987

By R. L. Beatty

The Bond International Gold (BIG) Colosseum Gold Mine was discovered in 1865, but no recorded production occurred until the 1930's, with production approximately 19k9 (615 oz) of gold. The mine was closed in 1942 as a non-essential industry during World War II. The property was acquired by Dallhold Resources in mid 1986. Construction of a 1.1 million Mt carbon-in-pulp, cyanide mill started in November, 1986, and was completed in September, 1987. Minable reserves are estimated to be 9,560,831Mt (10,539,000St), with an average grade of 2.1 g/Mt (0.062 oz/st). This paper describes a general view of process flow and plant equipment, as well as a look at the maintenance and operational problems encountered in the fist six months of operation.

Jan 1, 1989

By William Watson

Coal production in the United States in 2010 increased to a level of 960 Mt (1,085.3 million st), according to preliminary data from the U.S. Energy Information Administration (EIA), an increase of 1 percent, or 9.43 Mt (10.4 million st) above the 2009 level of 975 Mt (1,074.9 million st) (Table 1). In 2010, U.S. coal consumption increased in all sectors except commercial and institutional, while total coal stocks fell slightly for the year. Coal consumption in the electric power sector in 2010 was higher by 4.5 percent, while coking coal consumption increased by 37.9 percent and the other industrial sector increased by 7.1 percent. The commercial and institutional sector, which prior to 2008 had been called the ?residential and commercial? sector and is the smallest of all the coal-consuming sectors, declined by 3.1 percent in 2010. (Note: All percentage change calculations are done at the short tons level.) U.S. coal exports increased from the 2009 levels due mostly to higher demand for metallurgical coal, while coal imports decreased for a third year in a row. The increase in coal consumption for electric generation during the year was the consequence of improving domestic economic conditions further bolstered by a cold winter and warm summer in important coal-consuming regions. Preliminary data show that total generation in the electric power sector (electric utilities and independent power producers, including useful thermal output) in the U.S. increased in 2010 by 4.2 percent. Coal-based generation also increased, resulting in a 38-Mt (42-million st) increase in coal consumed in the electric power sector. Coal use in all the sectors other than the electric power sector increased by 13.9 percent to a level of 65.9 Mt (72.7 million st).

Jan 1, 2011

By K. C. Liddell

Chemical processing research made notable advances during 1986. This overview is necessarily brief. It is intended, though, to give the flavor of selected areas where recent activity has been high. At the 1986 Annual Meeting of The Metallurgical Society, a symposium was held on hydrometallurgical reactor design and kinetics. R. G. Bautista, R. J. Wesely, and G. W. Warren edited the resultant proceedings volume. This is the first available volume on this subject. It includes 26 papers on fundamental kinetic studies, modeling of reaction kinetics and reactor performance, and pilot and production-scale operations. Metals discussed include silver, gold, platinum, titanium, cobalt, nickel, zinc, copper, manganese, and iron. Agitated tanks, heaps, dumps, pachucas, pressure auto-claves, electrolytic reactors, liquid membranes, and fluidized beds are among the reactor types covered. An international symposium dealt with iron control in hydrometallurgy. It was sponsored by the Metallurgical Society of CIM, the Institution of Mining and Metallurgy, The Metallurgical Society of AIME, and Gesellschaft Deutscher Metallhutten and Bergleute. J. E. Dutrizac and A. J. Monhemius edited a proceedings volume containing 41 papers. These covered many areas of process chemistry, solvent extraction, precipitation, treatment of pickle liquors, impurities, residues and the environment, and process selection. Precious metals Concerning precious metals processing, the US Bureau of Mines IC 9059 is noteworthy. It includes papers presented at a USBM briefing at the Western Mining Conference in Denver. Topics covered included ion exchange, staged heap-leaching direct-electrowinning, and mercury precipitation during cyanide leaching of gold ores. Also dealt with were carbonaceous gold ores, carbon adsorption and desorption, heap leaching, the carbon-in-pulp process, and precious metals recovery from electronic scrap and solder. R. C. Sandberg and J. L. Huiatt (Journal of Metals, June 1986, and USBM RI 9022) developed a method to recover silver, gold, and lead from a complex sulfide using ferric chloride, thiourea, and brine-leach solutions. Gallium Gallium recovery was also a subject of considerable interest. Much information on this topic is proprietary. But two papers described solvent extraction of gallium. V. P. Judin and R. G. Bautista (Metallurgical Transactions B, 1986) developed an equilibrium model to separate gallium chloride from aluminum chloride. Tributyl phosphate was the extractant studied. T. Sato and H. Oishi (Hydrometallurgy, 16, 1986) investigated gallium extraction from sodium hydroxide by using Kelex 100. Data were given on the equilibrium distribution and the extraction kinetics. Galena Interest continues in recovering lead by hydrometallurgical processing of galena. J. E. Dutrizac; S. H. Kim, H. Henein, and G. W. Warren; and M. C. Fuerstenau et al. (Metallurgical Transactions B, 1986) all investigated leaching of PbS by ferric chloride. Dutrizac reported parabolic kinetics and reaction control by outward diffusion of lead through a porous layer of elemental sulfur. Kim, Henein, and Warren, however, reported that two leaching reactions occur. A nonoxidative reaction produces H2S and there is an oxidative leaching by ferric ion. Acid dissolution was found to predominate when the acid activity to ferric activity ratio is high. Surface chemical reaction was reported to be controlling. Fuerstenau et al. also reported that mass transfer through the sulfur layer is rate determining. But they believe that the species transferred are chloro complexes of ferric ion. In other work on lead, A. Y. Lee, A. M. Wethington, and E. R. Cole, Jr. of the USBM described an environmentally acceptable hydrometallurgical alternative to the smelting of lead concentrates (RI 9055). Uranium In spite of depressed prices, much work has been reported on uranium processing. R. G. L. McCready, D. Wadden, and A. March-bank (Hydrometallurgy 17, 1986) described the nutrient needs for in-place leaching by T. ferrooxidans. They reported the optimal conditions for uranium solubilization. L. E. Eary, H. L. Barnes, and L. M. Cathles (Metallurgical Transactions B, 1986) carried out an experimental and modeling study of uraninite dissolution. They concluded that ferric ion preferentially leaches uraninite in pyritic ores. Hydrogen peroxide was found to be less selective. A. Vuorinen, P. Hiltunen, and O. H. Tuovinen (Hydrometallurgy 15, 1986) studied redox and precipitation reactions of iron and uranium in leach liquors. P. T. Chiang (Hydrometallurgy 17, 1986) reported on the effect of uranium loading in the DEPA-TOPO process to separate uranium and iron from wet-process phosphoric acid. F. J. Hurst (Hydrometallurgy 16, 1986) conducted a fundamental study of the separation of uranium from phosphoric acid by DEPA and TOPO. D. E. Chia and W. C. Cooper (Hydrometallurgy 16, 1986) reported on bench and pilot scale work to recover uranium by the HIMIX process. They reported that the acid consumption to produce an eluate suitable for yellowcake

Jan 5, 1987

By D. F. Lillie

The milling plant of Bicroft Uranium Mines Limited has been in operation for a period of 2 years. During this time many changes, both physical and chemical, have been made in an effort to improve plant efficiency. Investigational work has been facilitated greatly by the fact that the Bicroft mill consists to a large extent of two parallel circuits; however, the number of inter-related variables is great, and it is difficult to measure, let alone control, all of them simultaneously. The Bicroft mill staff feel that much progress has been made, but there is much yet to be done and many questions still to be answered. The property of Bicroft Uranium Mines is located in Cardiff Township in the Highlands of Haliburton County, southeastern Ontario, some 140 miles northeast of the city of Toronto and 70 miles due north of Lake Ontario. The ore-bearing dikes are classed as mafic, non-segregated, syenite and granite pegmatites. Microline, microline perthite, quartz, uraninite, uranotharite, fluorite, titanite, cyrolite, calcite, pyrite, pyrrhotite and molybdenite as important accessories. The mill was designed for a nominal capacity of 1,000 tpd with all sections except grinding, leaching and clarification capable of 1,500 tons. Feed was entered to one mill unit on October 29, 1956 and to the second parallel line on November 13. The total feed rate was at 1,000 tpd on November 15. The first precipitate was produced on November 12. By August 1958 the average entry had been increased to 1,410 tpd without the installation of any additional major equipment. The overall recovery has been greater than 93,0 pet do several months, but a recovery of 92.5 pct appears to be most economical. Operating time for 8 months in 1958 has been 97.1 pet although all grinding mills were relined during this period.

Jan 1, 1959

By Jeff Hussey

Since 1955, Gaspe Mines has produced 150 Mt (165 million st) of ore averaging 0.9% copper. The ore was mined from two open pits and from underground workings. In 1994, four exploration drill holes discovered a buried porphyry/skarn copper-molybdenum (Cu-Mo) deposit at depths of between 1,000 and 1,700 m (3,280 and 5,580 ft) below the top of Porphyry Mountain. The mineralization and alteration associated with the Porphyry Mountain quartz monzonite does not outcrop. Prior to 1997, about 19,500 m (64,000 ft) of drilling in 12 drill holes had outlined 200 Mt (220 million st) grading 0.73% Cu and 0.08% Mo. Mineralization occurs as veins and disseminations of chalcopyrite and molybdenite.

Jan 1, 1998

By M. Mathiba

High CO2 fluxes associated with reclaimed coal mine spoils pose an environmental hazard where such sites are used for building. CO2 concentrations in excess of 25% have been reported in homes constructed on or down gradient of such spoils. The objective of this study was to investigate the effect of a mitigation trench used at one mine site on CO2 fluxes across the trench using chamber?based trace gas measurement. LI8100 automated CO2 flux system was used to measure CO2 fluxes from 32 sample points located across a deep mitigation trench on June 8 and 9. The maximum recorded fluxes were 7.29 and 8.18 µmol/m2/s, with averages of 4.15 and 5.50 µmol/m2/s, up and down-dip respectively. These preliminary results show CO2 flux is higher up-hill than down-hill at 95% significant level. This work shows that flux measurement can be used to investigate CO2 migration on reclaimed mine land. Further work is necessary to allow geostatistical modeling of the spatial variability of CO2 fluxes over reclaimed coal mine spoils. Knowledge gained would assist mining practitioners and regulators in reclamation planning for envisaged post-mining land uses.

Jan 1, 2010

By M. H. H. Mahmoud, T. A. El-Barbary, E. A. Abdel Aal, I. A. Ibrahim, A. K. Ismail

Fertilizer-grade phosphoric acid with a P2O5 concentration of 50% should contain less than 1.5% Fe2O3, otherwise, the P2O5 water solubility of the fertilizers prepared from high-iron acids will be low. High-iron Egyptian phosphate concentrates fromAbu-Tartur and Sebaiya- West were leached with sulfuric acid and recycled phosphoric acid to produce 50%-P2O5 phosphoric acids containing 4.0% and 2.7% Fe2O3, respectively. The high iron contents were reduced to the industrially accepted limit of 1.5% Fe2O3 by precipitation with potassium sulfate. The factors affecting the effi¬ciency of the precipitation process, e.g., temperature, time and K2O/Fe2O3 ratio, were systematically studied, and the optimum conditions were achieved. The clarification of the prepared phosphoric acid from iron-containing sludges was also investigated by varying the time and temperature. The utilization of the separated sludges as fertilizer and recovery of P205 by pulping in water were also evaluated.

Jan 1, 2000

By P. D. Rao, D. E. Walsh

A pilot-scale study was performed to evaluate the potential for gravity concentration of gold from -13 mm (-1/2 in.) crushed ore from Ester Dome, AK. 9.04 t (9.94 st) of crushed ore were sampled and processed through a Pan-American jig-Wilfley table pilot plant. Plant products were sampled for assay and material balance calculations. Gold recovery through one stage of jigging and two stages of tabling was 17.4% at a concentration ratio of 615:1. Concentrate grade was 766.3 g/t (22.4 oz/st). The feed grade to the plant was calculated at 1.71 g/t (0.05 oz/st). Jig tailings, primary table tailings and secondary table tailings grades were 1.44 g/t (0.042 oz/st), 1.30 g/t (0.038 oz/st), and 5.47 g/t (0.16 oz/st, respectively. Significant coarse, free-milling gold values were shown to remain in the, jig tailings. Concentrate mineralogy is discussed.

Jan 1, 1992

By John Barber

The Deep Ore Zone (DOZ) Mine is PT Freeport Indonesia's (PTFI) newest operation. Production at the mine began in September 2000. The DOZ Mine produces 25 kt/d (27,500 stpd) using mechanized block cave mining methods. The DOZ Mine has several major improvements when compared with previous block cave mines in the district. It has an independent and improved ventilation system, a flexible and efficient truck haulage system, and a more reliable crushing and conveying system. PTFI's copper and gold mining complex is located in the Ertsberg mining district in Irian Jaya, Indonesia (Fig. 1). The Ertsberg district is located in the Sudirman Mountains at elevations from 3,000 to 4,500 m (9,800 to 14,700 ft) above sea level. The topography is extremely rugged and rainfall in the mine area averages 3,000 mm/a (118 in./year). Freeport began production in the district in 1972 when the mill began processing ore from the Ertsberg openpit. Underground mining began in 1980 when the Gunung Bijih Timur-Ertsberg East (GBT) Mine was brought online using block caving methods. In 1991, the GBT Mine reached a maximum production rate of 28 kt/d (31,000 stpd). Ore from the mine was exhausted in 1994. The Intermediate Ore Zone (IOZ) was brought into production in 1994. It, too, used block caving methods and had a design rate of 10 kt/d (11,000 stpd). The IOZ is currently producing at a rate of 21 kt/ d (23,000 stpd). The DOZ was discovered in the mid-1980s by deep drilling from the GBT. The GBT, IOZ and DOZ deposits are all situated in the Ertsberg East Skarn System (EESS). In 1993, the first of several studies was completed that indicated that portions of the DOZ Mine could be economically mined using block caving methods. Limited development was started in 1995. Full-scale development and construction began in late 1997. The crusher and conveying system began operating during the first quarter 2000. The first of two main exhaust shafts has been com¬missioned, along with three of the four new main fans. The second exhaust shaft and fourth main fan will be in service by the end of 2001. Undercutting and production began in September 2000. The DOZ Mine will reach its design production rate of 25 kt/d (27,500 stpd) in 2003. A study is under way to determine if an ultimate production rate of 35 kt/d (38,500 stpd) is feasible.

Jan 1, 2001

By Yao Yevenyo Ziggah, Victor Amoako Temeng, Solomon Evans Kweku Koomson

Tyre hours/life deficit is a major operational challenge facing the mining industry which adversely affects materials production and costs. An accurate forecast of the tyre life is key in addressing this menace. This study for the first time employed the hybrid intelligent technique by utilising three metaheuristic optimisation algorithms, including particle swarm optimisation (PSO), genetic algorithm (GA), and whale optimisation algorithm (WOA), as trainers for the parametric weights and biases to optimise multilayer perceptron neural network (MLPNN) for enhancing prediction of on-site dump truck tyre life in the mine. Four hybrid models known as PSO-MLPNN, WOA-MLPNN, GA-MLPNN, and BP-MLPNN were developed using a total of 157 tyre dataset records obtained from a surface mine in Ghana. In assessing the prediction performances for the models developed, five statistical performance metrics of variance accounted for (VAF), Nash–Sutcliffe efficiency index (NASH), coefficient of determination (r2), mean absolute percentage error (MAPE), and correlation coefficient (r) were utilised. Moreover, ranking, uncertainty analysis and Bayesian information criterion (BIC) techniques were utilised to establish the most effective hybrid model, whereas sensitivity analysis was conducted on the input parameters. Results achieved showed that PSO-MLPNN was the best for prediction because it had the least MAPE value of 1.196% and relatively high values of VAF (99.642%), NASH (0.996), r2 (0.996), and r (0.998). Besides, PSO-MLPNN had the best selection criteria values of 6, 7.1725, and 444.834 for the ranking, uncertainty analysis and BIC respectively. Hence, PSO-MLPNN is recommended for the prediction of on-site dump truck tyre life for the studied mine.

Mar 11, 2024

By R. L. De G. Solms

INDEX 1. INTRODUCTION 2. FLOODING 3. THE PLUG 4. THE FIRST ATTEMPT 5. THE SECOND ATTEMPT 6. PUMPING 7. TIGHTENING OF THE PLUG 8. SHAFT REPAIRS 9. POST-MORTEM 10. CONCLUSION

Jan 1, 1987

By CHRISTOPHER D. HEBERT, SCOTT W. HOFFMAN

All Rights Reserved. Printed in the United States of America. Information contained in this work has been obtained by SME, Inc. from sources believed to be reliable. However, neither SME nor its authors guarantee the accuracy or completeness of any information published herein, and neither SME nor its authors shall be responsible for any errors, omissions, or damages arising out of use of this information. This work is published with the understanding that SME and its authors are supplying information but are not attempting to render engineering or other professional services. No part of this publication may be reproduced, stored in a retrieval system, or transmitted in any form or by any means, electronic, mechanical, photocopying, recording, or otherwise, without the prior written permission of the publisher. Any statement or views presented here are those of individual authors and are not necessarily those of the SME. The mention of trade names for commercial products does not imply the approval or endorsement of SME.

Dec 1, 2019

By Jameel Menashi

Aqueous solutions of Na2Cr04 can be reduced with CO to produce hydrous Cr203 and NaHCO3/Na2CO3. CH3OH, HCOO and HCHO in the presence of CO2 also reduce Cr 6+ and yield similar products. It is shown that the rate of Cr6+ reduction can be represented by the expression -d[Cr6+]/dt=k[Cr6+)1/2[R] where [R] is the aqueous phase concentration of CH3OH, HCOO- or CO. With these three reductants, the activation energies for the reduction processes are comparable and are of the order of 100 kJ/mole. Rapid reduction rates are attained at temperatures in excess of 220°C and, with these three reductants, a green acid-insoluble readily filterable hydrous Cr203 is formed. With HCHO significant reduction of Cr6+ occurs at about 80-100°C and an acid-soluble violet modification of hydrous Cr2O3 is formed. The reduction process is reasonably selective: solutions of Na2CrO4 contaminated with Mo6+ and W6+ can be reduced to yield a Cr2O3 product with significantly reduced contaminant levels. These findings are utilized as a basis for a process for recovering chromium from secondary source materials.

Jan 1, 1984

By Marcelo de Oliveira, Antoine Allanore, Dennis P. Chen, Davide Ciceri

Potassium fertilizers (potash) are mined in a handful of countries in the Northern Hemisphere. Potash is either expensive or unavailable to the farmers of the Southern Hemisphere. Alkaline alteration of K-feldspar rock (ultrapotassic syenite) in autoclave in the presence of Ca2+ generates a mixture of K-bearing calcium-aluminum-silicate-hydrate (C-A-S-H) minerals that potentially can be used as an alternative local fertilizer. This work investigates the effect of processing temperature (200 °C, 220 °C, and 230 °C) and time (from 0.5 to 3.0 h) on pH, mineralogy, and elemental availability of the C-A-S-H mineral mixture. Increasing temperature results in an increased conversion of K-feldspar and increased content of both the mineral tobermorite (Ca5Si6O16(OH)2·4H2O) and amorphous phase in the resulting C-A-S-H mixture; increasing time results in an increased conversion of α-dicalcium silicate hydrate into tobermorite. The amount of K leached from the mixture is relatively constant across all processing conditions. Hydrothermal alteration in autoclave of K-feldspar-bearing rocks such as ultrapotassic syenite permits the synthesis of a potential potassium fertilizer which alkalinity and mineralogy can be tailored according to soil requirements.

May 22, 2020

By Sandeep Kumar Jeripotula, Govinda Raj Mandela, Aruna Manglapady

Dozer operators are frequently exposed to whole-body vibration (WBV) during the execution of their work. Occupational exposure to WBV in Indian surface coal mines was evaluated by measuring vibration intensity and duration of exposure. A triaxial accelerometer was placed on the operator seat surface for taking the readings. Based on frequency-weighted root mean square acceleration equivalent to 8-hr shift duration, i.e., (A(8)) all dozer operators have exceeded an Exposure Action Value (EAV) of 0.5 m/s2 , and 90% of dozers did not exceed Exposure Limit Value (ELV) of 1.15 m/s2 . Based on Vibration Dose Value (VDV (8)), all dozer operators have exceeded Exposure Limit Value (EAV) of 9.1 m/s1.75, but no dozer operators have exceeded Exposure Limit Value (ELV) of 21 m/s1.75. Further, an epidemiological study was performed for identifying the extent of musculoskeletal disorders (MSDs) among dozer operators. For the detailed study, 42 dozer operators and 22 controls were selected from 2 surface coal mines. The control group was not exposed to WBV. It was seen from the cross-sectional study that pain in the lower back was predominantly higher (83.33%) in the exposed group when compared with the control group (31.81%). Likewise, pain in the neck (47.61%), shoulder (42.85%), knees (42.85%), and ankle (11.90%) was higher in the exposed group than that of the control group (22.71%, 0%, 45.45%, and 4.54%). A significant observation among the exposed group was that there was degradation in the quality of life. The outcome of the study would assist in monitoring and mitigation of machinery-induced vibration diseases (MIVD) in India and generally applicable to most of the mechanized mines as well. However, comprehensive studies are still needed to enunciate the magnitude extent.

By Yoginder P. Chugh

The control of soils has been found to be one of the highest priority factors influencing the return of reclaimed prime farmland to production. The compaction of soil is primarily a result of the movement of high ground pressure wheel scrapers and dozers over soils during the soils replacement and regarding process and occurs to a maximum depth of about 1.2 m (4 ft) The deleterious effects of soil compaction can be controlled by deep soil loosening to a depth of 1.2 m (4 ft). Most farm tillage equipment can only loosen soils to a maximum depth of 0.33 m (1.0 ft). Deeper soil loosening equipment, in the 0.33-1.2 m (1.0-4.0 ft) range is currently avail- able in the U.S. and abroad which may have potential for application on reclaimed prime farmland. This study has appraised such available equipment and the results indicate that a few manufacturers in the U.S., West Germany, and England market static and vibratory soil loosening equipment which have potential for deeper soil loosening of reclaimed prime farmland.

Jan 1, 1985

By Daniel E. Walsh

A pilot scale study was performed to evaluate the potential for gravity concentration of gold from -13 mm (-1/2"), crushed ore from Ester Dome, Alaska. 9.04 tonnes (9.94 tons) of crushed ore were sampled and processed through a Pan-American jig -Wifley table pilot plant. Plant products were sampled for assay and material balance calculations. Gold recovery through one stage of jigging and two stages of tabling was 17.4% at a concentration ratio of 615:1. Concentrate grade was 766.3 g/tonne (22.4 oz/ton). The feed grade to the plant was calculated at 1.71 g/tonne (0.050z/ton). Jig tailings, primary table tailings and secondary table tailings grades were 1.44 g/tonne (0.042), 1.30 g/tonne (0.038), and 5.47 g/tonne (0.16 oz/ton) respectively. Significant, coarse, free milling gold values were shown to remain in the jig tailings. Concentrate mineralogy is discussed.

Jan 1, 1991