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By C. Dan Kealy

Recent antipollution legislation for the control of mine tailings and waste-water has required an increasing number of tailings ponds. Many of these new waste-retention ponds are being constructed in congested locations, or the older ponds have reached such a size that--if failure occurs--they pose a threat to human life and public property. Although the typical tailings-disposal embankment represents--to the casual-, observer or the economy-minded mine official--a rather ugly but benign and stable pile of waste, they are in fact an extremely sensitive structural mass, retained behind the frailest of dikes. Failure of the waste mass requires only a minor change in water retention or weather conditions, erosion or piping of the ground strata, shifts in load or slight earth tremors. When failures occur, the results can be disastrous. In the U.S., we have been fortunate that, so far, pond failures have not caused loss of life. Thus, failures in this country have attracted little public attention. Present tailings-pond construction methods used by the mining industry require the barest money investment, and little engineering effort. Unfortunately, this superficial approach may increase short-term profits, but the practice may be conducive to the construction of needlessly weak structures. In coming years many of these designs will probaly become corporate liabilities. A typical pond is portrayed in Figure IA. This particular pond in Arizona is 1 mile long and 3/4 mile wide. Illustrated in Figure lB is a representative pond in cross section. This paper in concerned with two basic features common to all tailings disposal: (1) location of the phreatic surface within a tailings-pond embankment because of its influence on embankment-slope stability, and (2) prediction of

Jan 1, 1971

By Bernhard C. Koch, Richard W. Hutchinson, Bernhard Free

The Tertiary Silverton caldera collapsed asymmetrically to the southwest 27.5 m.y. ago. Resurgent volcanic activity is expressed by a quartz-monzonite stock at the southern caldera margin and by the Red Mountain porphyry stock in the northwest of the caldera which were emplaced at 25 m.y. and 11 m.y. b.p., respectively. The apical Eureka graben formed between 27.5-22.5 m.y. ago and transects the Silverton caldera southwestward. In the northeast it terminates at the 22.5 m.y. Lake City caldera margin. This graben as well as caldera faults provide dominant structural control for base and precious metal deposits. The epithermal gold mineralization of the Gold King mine is proximal to the Red Mountain stock in the north central part of the caldera. Here, west-northwest striking faults related to the collapse of the caldera intersect northeast faults of the Eureka graben. Superimposed upon earlier base metal mineralization at 23 to 22.5 m.y. b.p., the 11 m.y. younger gold metallization was a distinct event at Gold King. Supergene enrichment of the hypogene precious metal assemblage resulted in a metal tenor of up to 3,016 g/t gold and 1,088 g/t silver. The deposit is hosted by quartz-latite tuffs, flows, and flow breccias of the Burns Formation. Veins, up to 5 m wide follow steeply southeast and northwest dipping structures of the Eureka graben. Subsidiary veins up to 2 m wide branch off the main lode and are termed "flat veins" because of their conformable strike but shallow dip of 45' to 50' toward the northwest. Meteoric water-dominated hydrothermal solutions generated by the Red Mountain stock ascended through intensely fractured, ferruginous

Jan 1, 1988

By R. S. Rickard, B. Ball

INTRODUCTION Pyrite responds to the flotation reagents normally used in sulfide flotation systems. For example, it is well known that xanthates are good collectors for pyrite, and cyanide, sulfide and hydroxides are depressants. The particular response of pyrite to these and other reagents is used to advantage in several areas. By careful control of depressant and collector additions, pyrite can be separated from other sulfides by flotation. This is of obvious importance in base metal sulfide flotation where pyrite depression is often required to produce a marketable grade concentrate. Flotation of pyrite has long been employed in gold ore processing, and it appears that pyrite flotation from coal may become important in the future (1). It is probably a direct consequence of this relative ease of separation, plus the low economic value of pyrite, that detailed investigations on the chemistry of its flotation and depression are relatively limited. Gaudin (2,3,4) and others, in the mid-1950's published data on the flotation of pyrite, but detailed or consistent mechanisms were lacking. In 1955, Sutherland and Wark's book (5) was published, in which a large amount of empirical information, especially earlier work of Wark and Cox, was presented. More recently, in 1968, significant contributions have been made by Fuerstenau et al. (6,7) and Majima and Takeda (El. These papers introduced electrochemical explanations of the effects observed with xanthate and cyanide ions.

Jan 1, 1976

By D. Belin

Treatment of dilute acid mine drainage and similarly polluted water is becoming increasingly complicated. The pressure of economics is now exacerbated by legislation which classifies as hazardous waste the precipitates from conventional lime treatment plants. Since acid mine drainage invariably contains sulfate ions, one solution is to use anaerobic Desulfovibrio type bacteria to reduce the sulfate into sulphide ions which could then be used to precipitate out the heavy metals. Results are presented for both batch flask cultures and continuous column tests. Compost was used as the bacterial substrate in the column test which simulated heaps. Conditions changed from aerobic to the required anaerobic deep in the columns. Polluted feeds (up to 700 mg/l zinc) gave effluent containing well below I mg/l zinc over extended periods. Direct transfer of sulphide ions through a stream from an isolated container of bacteria is a good alternative.

Jan 1, 1991

By J. Sulaiman Lubis

Indonesia is a developing country in Southeast Asia with a population of some 180 million people. It has a stable political environment and currently enjoys high rates of economic growth. Indonesia has abundant reserves of natural resources, ample for its own needs and for export. Energy resources include oil, gas and coal. The oil and gas industries are well established domestically and for export. In the last few years, though, coal industry development has been rapid. Coal resources from bituminous to lignite amount to more than 30 Gt (33 billion st). That is more than enough to meet domestic needs and provide significant amounts for export. Figure 1 shows coal basins in Indonesia. Coals in Indonesia are generally young steaming coals. They are very low ash and sulfur, high volatile, bituminous B and C. The Indonesian government's state coal company is called Perum Tambang Batubara. The company has actively promoted the development of its coal for domestic and export markets. In particular, the government has placed priority on non-oil exports of which coal is a major element. The government strives to honor all contractural commitments for the export of Indonesian coal from contractor, private and government mines. From 1987 through 1989, coal production has increased from about 3 Mt (3.3 million st) to 9 Mt (9.9 million st). Exports have grown from I Mt (1.1 million st) to 3 Mt (3.3 million st). These are significant increases for a small industry with limited infrastructure. But with growing investment and development, these production and export levels will increase substantially, and Indonesia will be a significant world supplier of coal in the 1990s.

Jan 1, 1990

By Virginia T. McLemore

The Black Hawk district in New Mexico hosts an unusual type of deposit: arsenide five-element veins. These deposits are Ag-Co-Ni-Bi-As-bearing carbonate veins with local concentrations of U, Pb, Zn, Sb, and other elements. They are high-grade silver deposits (1000s g/t Ag), but low tonnage (<1 Mt) and difficult to find. By 1881, claims were staked and a town developed. It is estimated that >$1,500,000 was produced in 1883–1893, when Ag prices dropped due to the 1893 Ag panic. Solid Silver Mining Co. produced $600,000 from the Black Hawk mine; one car load contained $28,000 worth of silver! The town of Black Hawk grew to 125 population by November 1883. By the end of 1893, the town was vacant; nothing remains. The district was idle until 1916, and the mines operated sporadically since because of lack of available capital, water issues, and difficulty in finding ore shoots. The area is a favorite for mineral collectors. Tungsten was found in the district in 1935 in quartz veins associated with pegmatites. Metal production from 1881–1960 is estimated as 3,000 lbs Cu, 1,000 oz Au, 1,286,000 oz Ag, and 4,000 lbs Pb. In addition, 10,542 short tons of 2.7–71% WO ore and 615 short tons of fluorspar ore have been produced from the district.

Feb 1, 2025

By Kirsten Corin, Belinda McFadzean, Cyril O’Connor, Jenny Wiese

As orebodies being mined globally are of ever-decreasing grades, the focus on the concentration process has moved increasingly to the challenges faced in management of the gangue minerals. This is especially so in the flotation of platinum group minerals (PGMs), where head grades are typically in the region of 4–7 g/t. Hence, in PGM operations, a typical recovery of about 80 percent requires depressing about 85 percent of the gangue minerals. In a flotation cell the main control parameters are reagent type and dosage, froth depth and air addition rate. This paper reviews research projects carried out over an extended period aimed at optimizing the depression of gangue minerals: both those that are naturally floatable and those, such as chromite, that are entrained.  This has involved comparing the two major depressants used in the industry — guar and carboxymethylcellulose (CMC) — in the depression of the relevant gangue minerals, as well as investigating their consequent effect on the behavior of the froth phase. Because most PGM flotation plants treat the ore as if all the PGMs are associated with sulfides, copper sulfate is widely used as an activator, but this can also lead to inadvertent activation of gangue minerals. Finally, chromite levels must be kept below 3–5 percent to avoid any negative effects on smelter performance, and thus it is important to reduce chromite recovery which mainly occurs due to entrainment through, for example, froth depth control.   

By Barry A. Hillman

The rapid evolution of computer and field sensor technology is now providing the ability to integrate mine and office opera­tions. Digital maps and mine plans as well as CAD and GIS can be transferred to ruggcdized field computers for development of real-time mine surveys and real-time GPS machine navigation and monitoring. Surveyors can insure that their surveys arc accu­rate before returning to the office. Engineers and surveyors are also able to upload field computer-based surveys to mine plan­ning software and directly to mine equipment including shovels, drills, and dozers providing the operator of the equipment with a real-time view of their work space. Production information can be transmitted through dispatch and/or centralized mine servers to update mine plans with mining progress and production data. Static mine models then become a dynamic element of opera­tions, providing greater use of information that emphasizes the spatial aspects of the mine by providing real-time maps, produc­tivity charts, and graphs. This information can be disseminated through various mine communication systems to provide opera­tors and management with real-time feedback as well as daily and weekly reports. Integrated systems now being used in exten­sive trials arc reporting productivity increases of 4 - 7% or more thereby providing rapid payback to the mine.

Jan 1, 1998

By R. Fedel-Moen, R. W. L. Kimber, B. D. Williams, S. R. Ragusa

The use o f cyanide to extract gold from ores is widespread. In the process of extracting gold, cyanide may form metal cyanide complexes with a variety of other metals present in the ore. This study was carried out to investigate the biodegradation of the Cu(I)-cyanide complex and tetracyanonickelate under varying pH and nutrient additions. Detailed studies using shake flask cultures and bioreactors showed that naturally occurring bacteria residing on gold-leach-pad ore uti¬lized the Cu(I)-cyanide complex (CuCN) and tetracyano-nickelate, Na2(NiCN4), as carbon and/or nitrogen sources. Ammonium-nitrogen (NH4+-N) and/or cyanate (CNO-) accumulated in solution as the metal cyanide complexes were degraded. The accumulation of cyanate as a product is a concern due to its toxicity. Degradation rates were increased if a growth substrate (peptone, 0.1%) was added to the medium or if bioreactors were primed with peptone before addition of the metal-cyanide complex. Tetracyanonickelate was degraded at pH 8 without added peptone, but it was not degraded at pH 10. The addition of peptone was required for degradation of the Cu(I)-cyanide complex at pH 8 and pH 10. At pH 8, the addition of peptone resulted in the accumulation of cyanate, whereas, at pH 10, cyanate was not detected in the cultures. Biodegradation of metal-cyanide complexes was faster (approximately 1.5 to ten times faster) in bioreactors than in shake flask cultures. The addition of peptone was not required for degradation of metal-cyanide complexes in bioreactors. However, priming the bioreaetor with peptone resulted in higher metal-cyanide degradation rates over multiple cycles.

Jan 1, 2000

By Leons Kovisars

Two-thirds of the world copper produced in 1979 incurred cash costs between $0.35 and $0.89 per lb, averaging $0.62 per lb. North American costs are higher at $0.48, $0.88, and $0.68 per lb, respectively. Variable cash costs are $0.20, $0.58, and $0.39 per lb for the world; $0.29, $0.63, and $0.46 per lb for North America. Cash costs are relevant for comparing existing operations and evaluating market pricing during periods of excess supply. Analysis of full costs of production (cash costs plus income taxes, profit, and non-cash items) becomes relevant in evaluating the incentive price required to develop new production plants. In mid-1979 this cost (incentive price) was about $1.10 per lb; it is expected to approach $2.00 per lb by 1985.

Jan 1, 1979

By S. P. Chong, Peter V. Avotins

The iron and steel industry continues to operate at depressed production levels due to the low level of demand for steel and foundry products. Crude steel production in the Western world dropped by 13.7% from 459.3 Gt (506 billion st) in 1981 to 396.4 Gt (437 billion st). The reduction was particularly severe in North America where the industry suffered a loss of 37.3% from 123.6 to 77.5 Gt (136 to 85 billion st). This reduced level of production severely affected iron ore and agglomerates output. Substantial curtailment was made in both sintering and pelletizing plants to meet reduced demand. The pressure on these plant operators shifted from quantity to quality. With energy cost becoming an ever increasing part of production costs, conservation was a major area of emphasis. Iron Making Sinter has continued to be the favored burden for blast furnaces overseas. As hot metal production declined and excess sintering capacity developed, many of the overseas operators that have been using pellets found it advantageous on hot metal production cost to increase the ratio of sinter at the expense of pellets. This was the case despite the depressed price level of pellets in the world market. In North America where pellets have been the predominant burden material-mainly because it is necessary to pelletize the fine grain iron ores available domestically-there was also increased interest in sinter usage. Bethlehem Steel Corporation&apos;s 13.6-m (44.5-ft) furnace at Sparrows Point experimented with a burden with more than 70% of sinter during the second half of the year and indicated some favorable results as compared to its previous 50% super flux sinter and 50% acid pellet burden. The blast furnace operators continued to press for improved burden quality. In addition to the emphasis on the importance of burden chemistry and quality, uniformity, much emphasis was placed on improving reducibilitiy and high temperature metallurgical qualities. These burden properties were recognized to hold further promise in reducing fuels rates. Sintering With reduced consumption caused by the cutback in hot metal production, many sinter plants sat idle while others operated at greatly reduced rates. The deemphasis on production in many plants allowed them to experiment with various means to reduce operating costs. The most notable action was the introduction of low cost South American pellet feed fines into the sinter ore mix. These feed fines, which are finer than 150 µm (100 mesh) ore, were previously thought to be too fine for sintering. However, when the full production rate is not required the use of up to 40% in the ore mix was reported to be successful with the aid of burnt lime and adequate premixing and granulation of the sinter mix before sintering. The emphasis on quality and tighter cost control also has resulted in a further breakthrough in sintering technology. It was recognized that good feed preparation-including selection of compatible ores, good blending, good pseudo-particle formation, additive size control increased bed depth, and closed control of the sintering bed heat pattern-could allow the production of high quality sinters with much lower heat input and FeO level. Bed depth of more than 600 mm (23.6 in.) was practiced in a number of operations. Total heat input of better than 1.3 GJ/t (290,300 kcal per st) has been achieved. Kawasaki Steel&apos;s Chiba Works has also achieved a 5.3% SiO2 in the sinter with good quality. The low heat input approach allowed production

Jan 5, 1983

By Paul C. Lingo

Two recent antitrust cases, coming from widely separated areas, different industries and involving wholly different facts were decided by the Supreme Court on June 7, 1965. The decisions in there two cases may well prove to be the most important and far-reaching decisions in the field of labor law that have been handed down in the last 24 years. These cases, although quite dissimilar factually, had one common element. In both, the gravamen of the antitrust violation was bottomed on the terms of labor agreements negotiated by a multiemployer group. These cases are usually referred to as Pennington1 and Jewel Teat. Since a number of references will be made to the antitrust laws, it may be well to state briefly the applicable portions. Section 1 of the Sherman Antitrust Act makes "every contract, combination in the form of trust or otherwise, or conspiracy, in restraint of trade or commerce among the several states, or with foreign nations, is declared to be illegal...". Section 2 imposes criminal sanctions against "Every person who shall monopolize, or attempt to monopolize, or combine or conspire with any other person or persons, to monopolize any part of the trade or commerce among the several states...". Under Section 4 of the Clayton Act a person who sustains damages as a result of a violation of the anti¬trust laws is entitled to recover treble damages plus costs.

Jan 1, 1966

By E. B. Kroeger, Y. P. Chugh

The goal of this project is to develop, demonstrate, and document the beneficial use applications of a ponded F-fly ash as a road subbase material for the 2.1-mile Meredosia Industry Access Truck Route in Meredosia, Illinois. Approximately 60,000 to 70,000 cubic yards of ponded fly ash are proposed for use on this project without negative environmental impacts to surface water or groundwater. The County Highway Truck Route will consist of a 24-foot wide road built on a 0 to 7-foot thick compacted F-fly ash subbase. Laboratory testing indicates this fly ash will serve adequately as a road subbase, conforming to IDOT standard specifications. The economic considerations of using the Meredosia F-fly ash will also be developed so that a detailed cost-benefit analysis can be prepared for this project and for future projects. This project represents a two-year cooperative development and demonstration effort among the Office of Coal Development and Marketing (OCDM) of the Illinois Department of Commerce and Community Affairs (IDCCA), Southern Illinois University (SIU), Morgan County Highway Department (MCHD), Illinois Department of Transportation (IDOT), AmerenCIPS, and the Morgan County Regional Planning Commission. It is anticipated that monitoring activities will continue for a period of three (3) years after the completion of this project. The MCHD and IDOT will oversee all construction activities and long-term performance evaluation studies. Successful demonstration of large-volume beneficial use applications of F-fly ash at this site will significantly reduce management costs for power plants burning Illinois Coal. Furthermore, it will encourage similar utilization strategies at other sites in the State of Illinois.

Jan 1, 2002

By Wolfgang Pietsch

DR iron&apos;) is a generic term for metallized product that results from the direct or solid state reduction of iron ore. In direct reduction, iron oxides are reduced to metallic iron in the solid state without an intermediate melting step. MIDREX ® Iron is the DR iron material produced in the MIDREX ® Direct Reduction Process. Except for some thermal and mechanical degradation, the external physical form of as-reduced DR iron is the same as the feed material. Chemically, oxygen associated with iron is removed leaving a very porous structure. Figure 1 is the scanning electron micrograph (SEM) of the internal surface of a MIDREX Iron pellet obtained from a Swedish oxide pellet during reduction at 760 °C. Figure 2 is the SEM of a polished and etched cross-section through the same pellet showing the amount of intraparticle porosity. As a result of this high porosity, apparent density of DR iron is relatively low, typically a¬round 2 g/cm3, and specific surface is extremely high, typically around 1.0 m2/g or more. Both density and specific surface depend on the raw materi¬al and the operating conditions of the direct reduction plant. Normally, more than 97% of the available iron is reduced into metallic iron with the rest mainly in the form of wustite (FeO). During direct reduction, the gangue in the iron remains essentially unchanged. Removal of oxygen and the corre¬sponding reduction in weight result in a slightly higher percentage gangue in the product than in the raw material.

Jan 1, 1976

By W. D. Mayercheck, A. G. Mayton, J. R. Bartels

Since 1982, the US Bureau of Mines has investigated improvements in protection for operators of thin seam shuttle cars. Due to working height restrictions in thin seam mines (=1.22 m or 48 in.). Federal mine regulations require that canopy protection on shuttle cars be provided only in mines =1.07 m (42 in.). However, court obtained variances are granted to mines =1 .22 m (48 in). This paper presents ongoing research to design and evaluate a uniquely positioned, ergonomic compartment for a shuttle car that provides the operator with full canopy protection in working heights of 1.02 m (40 in.).

Jan 1, 1990

By Mark J. Zdunczyk

"Frac sand is industrial silica sand much like glass, foundry, filtration and blast sand. The sand must meet certain specifications set forth by the industry that consumes the product. The production and consumption of frac sand has experienced record growth in the last 10 years because of new oil and gas drilling, new horizontal well drilling, multi-stage fracturing and the stimulation of older wells that have little or no production. Frac sand is needed to “prop open” the fractures in the rock reservoirs to allow oil and gas to flow out more freely.According to the U.S. Geological Survey (USGS), the estimated production of frac sand in 2015 was 71 percent of the total production of industrial silica sand or 67.4 Mt (74.3 million st). This is a decrease of about 12 percent from 2014. However, the silica sand industry and other industry sources estimate only 50 Mt (55 million st) of frac sand were produced in 2015. In recent years both the USGS and industry studies have been very close in estimating the production of frac sand but with the recent downturn in frac sand production, perhaps, reporting to the USGS was not as accurate. These discrepancies may be that the USGS statistics are based on voluntary reporting by the industry and some companies report their approximate production, others may not. It is always difficult to obtain accurate production figures of any industrial mineral or construction aggregate operation as most companies tend to keep those figures somewhat confidential. It is certainly accurate to report that the production of frac sand had decreased since the last quarter of 2014 and through 2015. The oversupply of oil and natural gas has led to this decline in frac sand usage. Oversupply has idled drilling rigs and well completion activities. World Oil Magazine reports that 2015 was the worst year for rigs in a quarter century. The only wells being completed are those that have contract agreements to do so, others have been capped. Very few new wells are being drilled. Since about 98 percent of well completion is sand and water, the sand supply has caught up with the demand. There are some bright spots in the “oil patch” as the Permian Basin in Texas and the Woodford Shale in Oklahoma have kept pace somewhat with drilling and completion. It has been reported that the frac sand producers in the Brady-Voca, TX area are producing to nearly capacity. The northern frac sand producers like those in Wisconsin, Illinois and Minnesota are feeling the downturn because of the cost of shipping sand to these moderately active basins and plays. The “rule of thumb” is that freight is 2/3 and sand is 1/3 the cost."

Jul 1, 2016

By Dan Nilsson

INTRODUCTION There are many mine planning factors in sublevel caving, as in the other mining systems, which when varied can substantially alter mining costs and profit¬ability. In this chapter, the following four topics are addressed with regard to economic optimization of sub¬level caving: production planning, haulage level spacing, orepass spacing, and extraction cutoff. The costs used here are the right order of magnitude but since each mine is unique, they should be considered only as examples. The objective is to demonstrate the tech¬niques which can be used. In a real mine, the evalua¬tions must be done using actual costs and conditions. It is difficult to predict exactly the costs involved, and therefore it is valuable to perform a sensitivity analysis to evaluate the effect of making incorrect as¬sumptions. Since the mining industry is very capital intensive, the effect of the interest rate must also be studied. PRODUCTION PLAN FOR AN IRON ORE DEPOSIT Introduction The first and most important thing to do before de¬signing an underground mine is to establish a long-range production plan. Such a plan is necessary for all eco¬nomic evaluations and should provide information about the lifetime of the mine, how much ore and waste must be handled per year, how much development per year is required, etc. An example is given in the following section. Problem In an iron ore mine sublevel caving is used. The iron content is 42%, and the mine supplies a pelletizing plant with an annual capacity of 3 million t/a. The ore body is shown in Fig. 1. The length of the ore body is 1000 m and the width is 100 m. Each slice is 10 m high, and there are 10 m be¬tween crosscuts, each of which has an area of 20 m2. The density of the ore is 3.5 t/m3. The spacing between rings is 2 m, and the extraction is 100%. The iron con¬tent is 66% in the pellets and 6% in the tailings. The problem is to develop a detailed production plan. A typical sublevel caving sequence is shown in Fig. 2. Solution Amount of ore per meter of crosscut: 20 X 3.5 = 70 t. Number of crosscuts per slice= 100. Length of crosscuts per slice: 100 X 100 = 10 000. Ore from crosscuts per slice = 10 000 m X 20 m2 X 3.5 = 700 000 t. Number of sublevel caving rounds per slice: 10 000/2 = 5000. Area for each blast in sublevel caving: 10 X 10 - 20 = 80 m2. Amount of ore per blast: 80 m2 X 6 m X 3.5 = 560 t. Extraction: 100% (see extraction curve Fig. 12). Loaded ore per blast: 75% or 420 t. Loaded waste per blast: 25% or 140 t. Total: 560 t. Total amount of rock from each blast in the sublevel caving: Ore: 560 t of which 2 X 70 = 140 t from develop¬ment work and 420 t from sublevel caving. Waste: 140 t. Total: 700 t. The ore needed per year is 3 X (66-6)/42-6 5 mil¬lion t. Necessary number of blasts per year= 5,000,000/ 560 = 8929. Distance to develop per year = 8929 blasts per year X 2 m per blast 17 858 m. Total amount/year: Ore from development work 17 858 m x 70 t/m = 1.25 million t/a Ore from sublevel caving: 8929 blasts per year X 420 tons per blast = 3.75 million t/a 5.00 million t/a Waste rock dilution: 8929 blasts per year X 140 tons per blast 1.25 million t/a Total amount to hoist 6.25 million t/ a The amount of ore from development work will in¬crease a little if the horizontal drift is placed in the ore body and not in the footwall. In this example the ore loss in 20%, and the waste rock dilution is also 20%. After taking the ore loss into account, the lifetime for 100 m of the ore body will be:

Jan 1, 1982

By Larry D. Hartzog

For detailed information you are referred to "Heap Leach Solution Application at Rochester" by S. N. Dixon and A. L. Wilder which will be presented at the SME Meeting in Phoenix, January 1988.

Jan 1, 1998

By Babcock. C. O.

This report presents equations derived from experimental testing to convert borehole pressure cell data from vertical and horizontal cells to the biaxial ground stress changes. While hydraulic pressure cells in boreholes have been in extensive use since 1957, methods of data reduction have been slow to develop. The methods of Sellers (1970) and Bauer et al. (1985), respectively, indicate a linear relationship between cell pressure change and the rock stress change. Lu presented a theory for creep behavior (Babcock et al., 1982). In the present work, in laboratory testing, the logarithm of the cell pressure was related to the stress change in seven elastic materials for a range of Young&apos;s moduli of 3.45 to 85.5 GPa (0.5 to 12.4 x 106 psi). The gage response depends upon the setting pressure of the gage, the gage pressure at any time, the rock properties, and the cross-sensitivity of the gage to transverse loading. The setting pressure of the cell should be proportional to the Young&apos;s modulus of the rock, the absolute stress can be determined by overcoring in strong rock, and the in situ Poisson&apos;s ratio can be estimated in some cases from the cell responses.

Jan 1, 1986

By I. H. Warren

Hydrometallurgy has a well-known and established role in the production of manganese metal (1) and manganese compounds, including battery grade manganese dioxide (2). Also, several leaching processes have been developed and applied at various times for the recovery of manganese from low grade ores (3). Recently, interest in the hydrometallurgy of manganese has been stimulated by the problems presented in the processing of sea nodules (4). Fumes and dusts from the ferro-manganese furnaces also offer a potential field for the application of manganese leaching techniques (5). In treating low grade manganese ores a major requirement of the leaching process is that it should dissolve manganese from a mixture of manganese and iron oxides. The most selective reagent for this purpose is ammonium carbamate applied after preliminary pyrometallurgical reduction of the ore (6). Since soluble carbamate complexes of the postulated type (7) are formed by both manganous and ferrous, but by neither manganic nor ferric, ions, selectivety of manganese leaching is achieved by control of the reduction operation to yield, essentially MnO and Fe304? A potential improvement to this two step process would be to simultaneously reduce and selectively leach the manganese in one hydrometallurgical operation. This possibility has been experimentally studied and the results are described below. During the preliminary study of the stability of solutions of ammonium, ferrous and manganous carbamates at elevated temperatures, it was observed, contrary to the observation of Welsh (8) at lower temperatures, that

Jan 1, 1971