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By A. L. Williamson, G. Savci

Hydrologic assessments of existing and proposed waste rock stockpiles (WRS) are typically required for application of operating and closure permits at mine sites. The results of these assessments can be used to determine the applicability of the state groundwater permit program requirements to the operation and closure of these stockpiles. The objective of this hydrologic assessment was to ascertain whether or not the WRS at the Ajo Mine have a reasonable probability of discharging potential pollutants to the underlying aquifer under the operational mine conditions. The hydrologic assessment was performed based on an evaluation of climatic data, prediction of the net infiltration rate through the WRS and prediction of additional hydrologic factors such as the time required to deplete available moisture storage in WRS. The predicted net infiltration rates ranging from 0.13 inches/year (1.1x10-8 cm/s) to 0.24 inches/year (1.9x10-8 cm/s) are below the threshold value of less than 10-7 cm/s recommended by Hutchison and Ellison (1992) as being characteristic of conditions that produce no or negligible quantities of seepage. In addition, the total time required to deplete available moisture storage within the Ajo Mine WRS and for any potential seepage from the WRS to reach the underlying aquifer is estimated to range from 260 years to 1,950 years. Collectively, the climatological and geotechnical investigations and the hydrologic predictions demonstrate that the WRS at the Ajo Mine hydrologically do not have a reasonable probability of discharging pollutants to the underlying aquifer. As a result of this prediction, the stockpile will not be regulated under the state groundwater permit program and operational and reclamation requirements will be limited to managing surface water runoff.

Jan 1, 2002

By J. O. Marsden

Parts I and II in this series of papers presented the chemistry of medium-temperature pressure leaching of copper sulfide concentrates and reviewed the metallurgical development of a process to effectively recover copper using medium-temperature pressure leaching and direct electrowinning (DEW). Key features of the process were the use of superfine grinding of the concentrate (to a P98 of 12 to 15 µm and P80 of 6 to 7 µm), pressure leaching at 160ºC and 1,380 kPa (200 psi) oxygen overpressure, DEW to produce LME Grade A cathode, recycling of the lean electrolyte to the pressure leaching step and treatment of a lean electrolyte bleed by solution extraction (SX). Recycling of the lean electrolyte provided all of the acid required to effectively dissolve copper and convert approximately 65% to 70% of the sulfide sulfur to elemental sulfur, resulting in an ?acid-autogenous? process referred to as the ?MT-DEW-SX? process. This paper (Part III) discusses the steps taken to develop and complete a large-scale demonstration of the process at Bagdad, Arizona, in 2005.

Jan 1, 2007

By E. W. Thorton

"This paper is intended to show that grinding media can effect flotation results, and in some instances the results are vastly different. I do not suggest that autogenous grinding is beneficial in the separation of all complex base metal ores, in fact I know of a case where autogenous grinding would be disastrous to any selective flotation. I do however contend that prospective flow sheets and reagent balances of any complex base metal ores should be tried with both types of grinding. I feel fortunate that the paper to be presented later in this meeting by Rene Klymouski will, I think, compliment and perhaps explain some of the results of my research.This paper will be in four parts. The first part deals with a literature survey on the subject. Surprisingly enough, very little information has been printed on the subject. The second part deals with results obtained in bench scale tests on an ore from Chester Mines Limited. It was in these tests that we realized that steel grinding was giving us serious problems. So bench scale tests are shown giving results with a stainless steel grind, a mild steel grind and a grind in a ceramic mill with ceramic billets.The third part of the paper deals with results obtained in the Mines Branch Pilot Plant, when again we ground the ore with three different types of grinding media. Incidentally, this has been the only pilot plant test made at the Mines Branch in which flotation products were made on both steel grinding and autogenous grinding on the same ore. The fourth part deals with two other New Brunswick ores and is intended to show that changing the grinding media effects ores other than that of Chester Mines Limited."

Jan 1, 1973

By T. A. Kirk

"The Linatex Dewatering System is a compact, low capital and maintenance cost matched set of processing equipment suitable for solids recovery, dewatering and desliming sand-like materials in the range 3/8"" to 325 mesh ( 1). Since its introduction to North America in the early 1970s, various elements of the system have been successfully installed in a variety of applic11tions both in Canada and the U.S.A. Examples are chosen from the coal, fine aggregates, iron ore and industrial minerals industries.II. Elements of the Linatex Solids Recovery System(A) The Linatex Separator (See Fig.I)The Linatex Separator, which is the heart of the system, is a specialized type of hydrocyclone designed to give a self regulating rope discharge product. It has the familiar cyclone geometry of cylindrical and conical section, tangential inlet, apex and vortex finder. The apex is enclosed in a unique patented automatic underflow regulator consisting of a collapsible rubber sleeve, and the overflow line is a siphon-leg, with a controlled air bleed valve. The siphon leg is adjusted so as to balance the internal negative pressure acting as a closing force on the underflow regulator against the mass of slurry pressing down on it from the cone section of the Separator. With correct adjustment the underflow regulator will close drip-tight on a water only feed. As solids accumulate in the apex of the cone the weight of solids counterbalances the vacuum and cat1ses the underflow regulator to open and discharge classified material at a controlled density."

Jan 1, 1978

By Frederick Bradley Henderson

This study was initiated to investigate the occurrence of cinnabar in the oxidation zones of some mercury sulfide deposits. Seven mines or mine districts were visited and samples were taken from both the oxidation and primary zones of these deposits in order to compare their mineralogies and assess Vie chemical changes that took place during oxidation. As a result of the study of these specimens, several limiting factors are suggested for the chemical environments and processes active in both the deposition and oxidation of mercury sulfide deposits. The Mt. Jackson mine, near Guerneville, California (location map, fig. 1), was studied in some detail and forms the basis of this paper. The mine has not heretofore been described in detail in the literature. The Mt. Jackson mine is considered typical of mercury sulfide deposits occurring at the faulted boundaries of serpentinite bodies and enclosing sediments. Such mines are common in the California and Oregon Coast Ranges? Specimens collected by the author from the New Idria and New Almaden districts, the Mt. Diablo, the Culver Bear, and the Buena Vista mines in California, and the Opalite district (Cordero and Bretz, mines) on the Oregon-Nevada border were studied for comparison with the Mt. Jackson samples. The mine description, maps and sketches were composed from field maps kindly provided by Francis Fredericks, consulting geologist to the Sonoma Quick-silver Mine Inc. during the author's visit in 1962. Additional information was obtained from the company maps and minor corrections and interpretations were added by the author.

Jan 1, 1965

REFERENCES AND FOOTNOTES 1. U.S. Department of the Interior, Office of Surface Mining. Surface Coal Mining Reclamation: 10 Years of Progress: 1977-1987. Washington, D.C. August 3, 1987.48 pp. (For sale by the Superintendent of Documents, U. S. Government Printing Office, Washington D.C. 20402.) 2. The Rural Abandoned Mine Program (RAMP), authorized by Section 406 of SMCRA, provides for the control and prevention o f erosion and sediment damages from unreclaimed abandoned mine lands. 3. The Small Operator Assistance Program (SOAP), authorized by Section 401(b)(1) and 507(c) of SMCRA authorizes up to 10 percent of the fees collected for the Abandoned Mine Reclamation Fund to be used for technical assistance to help qualified small mine operators obtain techincal data needed for permit applications. Operators who produce fewer than 300,000 tons of coal per year are eligible for assistance. 4. OSM regulations 30 CFR 816.42 and 817.42 require water from areas distrubed by mining activities be in compliance with all applicable state and federal water-quality laws and effluent limitations for coal mining established by the U.S. Environmental Protection Agency. 5. Sourse: Abandoned Mine Lands Fee Collection System, Data from OSM-1 form. 6. Sourse: Abandoned Mine Lands Fee Collection System, Data from OSM-1 form. 7. Source: OSM Field Office Annual Reports. 8. Johnson, W. and G.C. Miller, Abandoned Coal-Mine Lands: nature, extent, and cost of reclamation. U. S. Bureau of Mines, 1979,29 pp. PHOTOGRAPHIC CREDITS Front cover, pp. i, 1, 2 bottom, 7 right, 10 bottom left, l 1 top, l l bottom right, 12 top, 13 bottom, 14 bottom right, 15,16 bottom, 17 top, 17 bottom right, 18 bottom, 19, 20 bottom, 21 top, 22, 23, 24 bottom, 25 bottom, 26 bottom left, 26 bottom right, 27, 28, 29 top, 30 bottom, 32 bottom right, 33 top, 34 top, 35 top left, 35 bottom left, 35 bottom right, 36 top, 37, top right, 37 bottom, 38 bottom, 39, 41, 42 top, 43 bottom, 44 bottom left, 44 bottom right, 45, 46, 47, and 53: Chuck Meyers, OSM. Page 2 top and center: Tony Fuhs. Page 3: John Sharkowicz, OSM. Page 6 left: North American Coal Corp. Pages 6 right, 7 left, l l bottom left, 13 top left, 13 top right, 17 bottom left, 21 bottom, 25 top left, 25 top right, 29 bottom, 30 center, 32 top, 32 bottom left, and 38 top: Bob Welch, OSM. Page 10 top left: Donald Stump, OSM. Pages 10 right, 14 top right, 14 center right, 18 top, and 44 top: Ron Hill, EPA. Page 12 bottom: Texas Utilities Co. Pages 16 top, 20 top, 24 top: Tennessee Valley Authority. Page 26 top right and 26 center left: Richard Dick, U.S. Bureau of Mines. Pages 30 top, 33 bottom left, 33 bottom right, 35 top right, 36 bottom, 37 top left, 40 top, 40 bottom, 42 bottom, 43 top left, and 43 top right: Jim Boyer, OSM. Page 31 top left, and 31 top right: Bruce Wage, Western Energy Co. Page 31 bottom: Rex Wilson, OSM. Page 34 bottom: West Virginia Division of Abandoned Mine Lands. Page 35 center: William Weist, Jr., OSM.

Jan 1, 1992

By G. H. Lind

Research into the safe and economic extraction of underground coal pillars in the Witbank and Highveld coalfields of South Africa highlighted that future practices will have to consider pertinent risks associated with this least safe of all underground coal mining methods conducted in South Africa. To this end, a computer-based program called A-PEP (Analysis for Pillar Extraction Potential) was developed to assess three aspects of this mining practice, viz. risks associated with the physical attributes, operational risk attributes and economic attributes. This intelligent tool can thus be used when planning to conduct underground pillar extraction to ensure that all pertinent risk elements and legislative requirements are considered to conduct this mining method safely before a submission is made to the Department of Minerals and Energy. Keywords: operational risk management, mine planning, underground coal mining, pillar extraction.

Jan 1, 2003

By C. Ogasawara

This document describes the direction and progress of the Bureau of Mines Oil Shale Mining Environmental Research Program as of January 31, 1979. It is aimed at focusing the Bureau's role in environmental research supporting oil shale development with respect to the activities of other involved government agencies and private industry. Environmental considerations in oil shale extraction, such as ecological disturbances, health and safety hazards, waste management, and socioeconomic impacts, are discussed along with the type of research known to be planned by organizations to ameliorate these problems.

Jan 1, 1979

By Hongyang Cao

The iron constituents selectively separated from the modified copper slags by magnetic separation and flotation.were studied. The results of magnetic separation for modified copper smelting slags show that the grinding size, current intensity and dispersant agent had great influence on grade, and recovery ratio of TFe in preparation concentrate; The results of flotation for magnetic concentrate show that the positive flotation of sodium oleate negatively upgrade the grade of magnetic concentrate, but the reverse flotation of lauryl amine and soluble starch are effective. The grade of TFe can reach 54% and the recovery ratio can reach about 90% in magnetic concentrate, when the grinding size is 45.8µm and the current intensity is 2.5A. The optimal effect is obtained by using lauryl amine as collector and soluble starch as depressant to reverse flotate magnetic concentrate, and the grade of TFe is upgraded 2.37%.

Jan 1, 2009

By L. Holappa

Ferroalloys like FeCr and FeMn are mostly produced in submerged electric arc furnaces. Typical of these processes is that the amount of slag is quite high. The role of slag for the process is thus essential and specific chemical, physical and electrical properties are required. Slags in FeCr process are Al2O3-MgO-SiO2 based with minor contents of CaO and CrOx. Slags in the FeMn process are principally similar, originating from silicate minerals but can have quite large composition range depending on the manganese ore and its gangue minerals. Characteristic of the FeMn process is that the slag is rich in manganese oxide up to above 30%. A common feature is that the ?ore oxides?, CrOx and MnOx in slag are not only divalent oxides but also contain some higher valent of Cr3+and Mn3+, respectively. In FeMn production conditions the slag consists mainly of Mn2+. This is a clear difference from CrOx containing slags in which both two and three valent coexist in slags in FeCr process conditions. Oxidation state of the slag depends on the ambient oxygen partial pressure and temperature in the examined process. The thermodynamics of slags in ferroalloys, production are discussed in terms of phase diagrams, activities and oxidation state of chromium and manganese oxides. Further, slag properties influencing reduction kinetics and furnace operation i.e. viscosity and electrical resistivity are discussed. Keywords: review, slags, ferroalloys, ferrochromium, ferromanganese, thermodynamics, physical properties

Jan 1, 2004

In redesigning the cyanide plant at Mount Torrens, in conjunction with Mr. L. W. Grayson, the manager, several minor alterations were introduced in order to have the operations under thorough control, and these having proved serviceable may be of sufficient interest to merit a description here. The lixiviation vats, though only 8 feet in diameter, were retained, but the pipes leading away from the bottom were reduced from two inches to half-an-inch in diameter, and had a fall to the zinc boxes of about 4 feet. As these small pipes were completely filled by the solutions fiowing away from the vats, they exerted a suction equal to their fall and thus hastened the rate of flow, so that tailings containing a larger proportion of slimes could be treated than was possible where the larger pipes were used, as, there not being sufficient fiow of solution to keep the latter filled, no suction could be obtained with them. It was found that with'most of the tailings treated a sufficient flow could be obtained from the half-inch pipes when the taps on, them were only one-third to half open. The lower end of each of these pipes was connected by means on indiarubber tubing with the sille hole of 11 in aspirator bottle, and to the mouth of this bottle a few feet of similar tubing was connected, so that the solutions coming through could be directed into either of the fixed tubes leading respectively to the waste water tank the strong solution zinc boxes or the weak solution zinc boxes.In working the vats, the following method was found to give good results. While filling the vats with tailings the taps were left open, so that the water might drain off into the waste water tank the surface of the tailings was carefully levelled. When the strong solution was run in at the top, the air contained in the interstices of the tailings was forced through the half-inch pipe at the bottom as a strong blast, instead of, as I believe js usual, escaping by means of channels formed by it through the tailings, and thus tending, it would appeal, to render the disposition of the tailings less even and to make.

By Peter Moser, A. Gaich, E. Grasedieck Zechmann

In The SMX Blast Controller is a novel tool for precisely controlling the geometry of a blast based on the technology of 3D imaging. It enables to derive precise information about the geometry of the face of a bench in a quarry from a pair of stereoscopic images taken from the bottom level of a bench with a calibrated standard digital camera. Apart from staking out the blast position with delimiters that mark the area to blast and taking then two images (freely), no further actions are required which allows performing the data acquisition without requiring any surveying skills. The 3D image generation and blast planning does not require special computer. It is estimated, that every blast foremen is able to work with the system after a one-day training as the system includes a high grade on automatism. The software package allows the blast foreman to design and stake out a blast, based on provided geometric values, such as burden, side spacing, hole inclination, sub-drilling, row number & distance and drilling pattern. The software connects these values with the 3D information acquired by the digital images. All the drill & blast information can be used also for blast documentation. The SMX Blast Controller overcomes also the problem of doing the back transfer from maps derived by surveying into nature as the resulting 3D image and all the information associated to it (hole position, etc.) can be correlated with the nature by visual comparison and the help of a single reference line, defined by two points. The system is designed to be used also in small to medium quarries. It can help, that the control of the geometry of the bench face and the precise design of the charge of the blast on a borehole per borehole basis will become an accepted daily routine.

Jan 1, 2006

By H. A. Domian, H. Biloni, G. F. Bolling

Jan 1, 1965

By J. Leinart

As is readily apparent, the reliability of the early-stage costs estimated for any mining project are critical to the viability of that project. An estimate too high might leave a potentially profitable deposit undeveloped or worse yet, available to a competitor. One too low might waste investment dollars in a project not yet economically ripe. There are two aspects critical to any estimate: 1. The cost data used to represent the values of the wages, salaries, supplies, and equipment. 2. The design used to determine the engineering parameters upon which the above estimated values are based. Given that all of the information contained in Mining Cost Service is provided by suppliers, equipment manufacturers and mine operators, one wonders why there is any question as to the reliability of cost estimates developed using this information. Given that it is all actual (as opposed to statistically derived) data, where do the problems with estimate reliability lie and what is the impact of the data on that reliability? The following presents a series of analyses designed to assess the reliability of the cost estimating process. Both the actual cost data and the engineering parameters upon which they are based are examined in an attempt to quantify and understand the importance of each. While several sources of cost information are available (Means, Richardson, Primedia?s Equipment Watch, and individual suppliers and manufacturers), the following rely solely on data contained in InfoMine USA, Inc?s, Mining Cost Service and in its Sherpa Mine Cost Estimating Software.

Feb 27, 2013

By Yushiro Hirai

Saganoseki smelter & refinery had operated two flash smelting furnaces from 1973, attaining 330,000 mtpy of copper production in 1989. In 1996, however, it carried out a "integration of two flash furnaces". That is, it doubled the feeding capacity to one flash furnace to shut down another furnace while maintaining the same production. Single flash furnace operation has successfully started in March 1996 accomplishing a great reduction of the total production cost. The production had reached to 350,000 mtpy in 1997, and further increased to 450,000 mtpy by executing the expansion program with single flash furnace in 1998. This paper introduces results of the single flash furnace operation and describes about an outline of the expansion.

Jan 1, 1999

By S. McCalpine, W. Muir, P. J. Downey

"Golden Bear Operating Company Ltd. (GBOC) was a joint venture between Chevron Minerals Ltd. and North American Metals Corp. that was formed for the development of the Golden Bear Project. The mine is located at Muddy Lake, Western British Columbia, approximately 80 km northwest of Telegraph Creek and 135 km west of Dease Lake (see Figure 1). Road access to the mine was developed by Golden Bear and consists of a radio controlled 155 km gravel road originating at highway 114.A feasibility study was completed in 1987 and project approval was finalized in 1988 with Golden Bear issuing a ""Request For Proposal"" for detailed engineering in May 1988. Golden Bear also commenced negotiating with vendors at this time for the supply of the two principal process equipment packages: the dry grinding and classification system and the whole ore roaster package.Process plant engineering and plant construction was completed in October 1989 with plant startup commencing in November 1989. Two months into the commissioning phase it became apparent that the plant, as designed, would not achieve the anticipated production targets.Kilborn Engineering Pacific were subsequently contracted by GBOC to review and modify the plant design, assist the operating personnel with plant commissioning and startup and provide operator training.This Paper describes some of the problems encountered during the startup and the plant modifications incorporated to achieve design parameters."

Jan 1, 1992

By Benkendorff P. N

Cement production in Australia approximates 6.5 million tonnes per year. Cement clinker is sourced from 15 plants which operate 28 kilns. Cement grinding is conducted at 14 of the clinker plants plus 7 grinding plants. Average annual capacity per kiln or plant is lower than in many overseas countries, but kilns introduced since 1970's have been of larger size more in keeping with size trends overseas, although still smaller than the large units overseas. Substantial economies in capital and labour costs can be achieved in large production units, particularly with dry process kilns using precalcination techniques in which part of the fuel is introduced in the static preheater section ahead of the kiln while the balance is added to the kiln in a conventional manner. Precalcination techniques have been introduced to the Australian cement industry but as cement transport costs form a major part of supplying diverse market areas, the size of kilns is likely to remain less than maximum owing to balance with distribution costs. The cement industry uses significant amounts of thermal and electrical energy which can be minimised by proper selection of equipment and technology. Labour costs are also significant and can be contained by automation and computer control of the process.

Jan 1, 1989

By Peter Radziszewski

"Typical mining operating costs can be divided between extraction (30 % to 70 %), comminution (30 % to 50 %) and separation (5% to 20% ). Cornminition process costs, representing an important contribution to mining operating costs, can themselves be divided roughly between energy (50 %) and linerkharge wear (50 %). Developing the ability of predicting mill wear will allow a more accurate estimation of wear costs. However, total media wear in a given ball mill grinding process is a product of three recognized wear mechanisms those being abrasion, corrosion and impact. Further, the contribution to total media wear of each of these wear mechanisms has not been well established. In this paper, a overview of abrasive, corrosive and impact wear mechanisms is made from which an empirical wear prediction procedure is proposed. Further, future directions in wear prediction research is outlined. INTRODUCTIONTypical mining operating costs can be divided between extraction (30 % to 70 %), comminution (30 % to 50 %) and separation (5 % to 20 %). Comminition process costs, representing an important contribution to mining operating costs, can themselves be divided roughly between energy (50 %) and lined-charge wear (50 %). With respect to wear, world wide steel consumption as grinding media is estimated at over 600 000 tomes per annum (Malghan, 1982). As such, developing the ability of predicting mill wear will allow a more accurate estimation of wear costs.Charge media wear prediction is actually estimated using the Bond abrasion test. From Bond abrasion index tests, it is possible to determine a wear rate using the following function:"

Jan 1, 2000

The area (330 sq. miles) shown on the map of the Brocks Creek District (Fig. 1), was mapped by the Northern Territory geological party of the Aerial, Geological and Geophysical Survey of Northern Australia. This work was done in two stages:-(i) In 1935, an area of 30 sq. miles around the Golden Dyke mine was mapped by P. S. Hossfeld, Senior Geologist, and A. H. Voisey, Assistant Geologist. The results were described in a detailed report (10).(ii) In 1939, an area of 300 sq. miles in the Brocks Creek district was mapped by C. J. Sullivan (Senior Assistant Geologist), and V. M. Cottle and S. N. Kiek (Assistant Geologists) under the general supervision of P. S. Hossfeld (Senior Geologist).An outline structural map, together with some geological notes was published (1), but that map does not show the amphibolites, and the relationship between ore occurrence and general geology is not emphasised. The publication of the detailed maps and reportwas postponed indefinitely as a result of war. The original mapping was based on air photographs with a scale of approximately 1 in. to 1,200 ft., and all detail which could be shown at that scale was mapped. Features recorded included rock types, details of structure, mineralisation, and surface workings, creeks and soil boundaries. Underground mapping was also carried out. The present map is much simplified from the original, but it is hoped that it shows sufficient information to reveal close relationship between structure, granite, sills of amphibolite and gold occurrence. About 50 per cent. of the area is covered by deep soil or alluvium, and in these portions the geology of the underlying rocks was not determined with accuracy, though in many places strike trends could be deduced from air photographs.Because the A.G. & G.S.N.A. had published the geological notes and map referred to above, and because the publication of the detailed report and maps is still postponed indefinitely, approval has been given by the Bureau of Mineral Resources, Geology and Geophysics (which took over the affairs of the A.G. & G.S.N.A. at the termination of its operations) to the preparation of the present paper so that information would be available to the public.The views expressed in this report are those of author. The results of the investigation suggest that in areas where granites are predominantly concordant with bedding, ore deposits may be found at approximately the same stratigraphic horizon. It is thus important to correlate beds even though their relationship may have been obscured by faulting and folding.

Jan 1, 1950

By K. S. Liddell

Ultra fine grinding is becoming increasingly used in the mining industry over recent years. The development of large-scale units for high volume grinding of low value materials and concentrates has opened up new areas of economical metal extraction. The ability of stirred vertical mills to produce very fine particles with high surface areas, is providing mine operators with novel extraction possibilities within gold concentrate regrind for Biox or leach feed, Activox autoclave feed preparation and lime slaking operations. The paper will review design, operational and maintenance aspects of ANI-Metprotech stirred vertical mills, comparing these to tower mills, horizontal bead mills and attrition mills.

Jan 1, 2001