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By S. Banisi, M. Pourkani, A. Kargar, D. Hamidi, M. Sarvi

"Flotation circuit of the Sarcheshmeh copper mine, which processes 40,000 t/d of ore at 0.9% Cu, consists of rougher, cleaning and recleaning stages. To study the stage recovery of copper, one of the eight identical flotation banks was selected for sampling. It was observed that 17% of copper goes to the tailings and the lowest copper recovery was for particles coarser than 105 microns. The third five-cell unit of this bank not only did not recover appreciable amount of copper but also decreased the concentrate grade. The grade of the last five cells was lower than the feed grade suggesting that mixing of the concentrates of the three units was not beneficial. To alleviate the problem, the concentrate of the first four-cell unit was directed to the cleaners feed box (i.e., by- passing the regrind mill). This increased the overall copper recovery by 1.7 %. Also rather than the current practice of adding the reagents only at the feed tank (i.e., 100-0-0%), various combinations of stage addition were examined. It was found that 75-25-0% addition regime could increase the total copper recovery by 1.3%. At the recleaning stage in order to increase the final copper concentrate grade, a froth washing system was used. To ease the maintenance of the system, a shower type washing system was installed. The optimum wash water superficial flowrate was found to be 0.06 cm/s. Continuos sampling of the recleaners with and without wash water over a period of two months showed an increase of 1.5% in the final copper concentrate grade at constant overall copper recovery.INTRODUCTIONThe Sarcheshmeh copper ore body that may rank as third or fourth largest in the world contains 1 billion tonnes averaging 0.9% copper and 0.035% molybdenum. It is located in southeastern Iran and has been processing 40 000 t/d of ore since 1982. The mine produces 100 000 tonnes of copper and 2200 tonnes of molybednite concentrate per year.After three stages of crushing, the ore with a F80 (80% passing size) of 12.7 mm is fed to eight parallel ball mills in closed circuit with cyclones to produce a product 70% finer than 75 microns. The concentrate of rougher flotation circuit is ground; rougher tails constitute the main component of the plant final tailings. A concentrate with an average grade of 31% Cu (chalcocite and chalcopyrite) is obtained after cleaning and re-cleaning stages. Total recovery of the plant varies between 83% and 87%, depending on operating conditions and ore type."

Jan 1, 2001

This paper describes the back analysis of the Pavoncelli tunnel, an hydraulic tunnel which was excavated between 1904 and 1915 in South of Italy and experienced severe damages during the 1980 Irpino-Lucano earthquake. A numerical simulation which is intended to study the behaviour of the tunnel during this seismic event is presented, also accounting for the swelling phenomenon that took place during and after construction.

May 1, 2009

The history of zinc production in Australia is outlined including attempts to make zinc electrolytically at Cockle Creek, N.S.W. in the last century, the use of horizontal retorts at Cockle Creek and Port Pirie, South Australia, production by electrolysis at Risdon, Tasmania and at Port Pirie using zinc oxide fumed from lead blast furnace slag and production by the Imperial Smelting process at Cockle Creek. Developments in the electrolytic process are described including those in roasting, acid manufacture, leaching, purification, electrolysis, melting and casting. The Imperial Smelting process technology is also outlined. In both processes global technology is reviewed, including important Australian contributions.

Jan 1, 1986

By M. J. Fairweather

"The development of means to maximize the selectivity between galena and sphalerite in lead zinc ores is extremely important to the lead-zinc industry. Two challenging studies recently undertaken by Cominco will be reviewed. The first required an unusually fine grind to achieve liberation between galena and sphalerite. The second contained partially altered galena which resulted in unusual early activation of sphalerite under normal flotation conditions. Means by which lead and zinc separation were optimized in these two ores will be discussed.There are two important factors in the separation of minerals by froth flotation: particle size distribution (grind) and pulp chemistry. Pulp chemistry includes a number of factors including pH, Eh, collector type and concentration, inorganic depressants, organic depressants and activators.Grind is important for two reasons. A sufficiently fine grind is required to optimize liberation. However, if the mineral dissemination is too fine, mineral selectivity in the flotation process is impaired. Due to the hydrodynamic constraints of conventional mechanical flotation cells, flotation works best over the single decade size range between 10 and 100 microns. The majority of new ores treated by Cominco until relatively recently were comparatively simple. The major factor in their treatment was defining optimum grind size. Accordingly, the focus within mineral processing at Cominco was on grind and supporting liberation studies. The relationships between grind, liberation and flotation response are intuitively easy to understand. An idealized relationship is shown in the first figure."

Jan 1, 1991

By A. A. Debrah

Artisanal and small-scale mining (ASM) plays a fundamental role in the national and local economies of resource-rich countries in Africa. As such, more effort must be geared towards fostering this type of economic activity to sustain livelihoods in rural communities. Across Africa, efforts to operationalize ASM based on a sustainable development framework have been quite tedious and often ineffective. This is because most of these frameworks lacked context and an understanding of the continuum from the artisanal through small-scale and up to the junior miner, grouped under the general term ?ASM?. Driven by the need to contextualize some of the definitional issues in this niche sector, and gain a clearer understanding of their peculiar challenges, this paper aims to contribute to the debate on ASM by considering two of Africa?s resource-rich countries ? Ghana and South Africa. Since South Africa has made greater inroads in protecting the interests of small-scale miners, we thus draw comparisons from both countries? approach to ASM, outline the lessons learnt, and provide recommendations regarding the ways forward.

Jan 1, 2014

By D. F. Malan

In deep level hard rock mining the rheological behaviour of the rock is important as time-dependent aseismic processes can lead to a stable dissipation of energy thus reducing violent outbursts. Underground closure measurements, observations of time-dependent fracture formation and siesmic data indicate that the rock in deep South African gold mines shows significant time-dependent behaviour. The objective of this work is to quantify the time-dependent rock behaviour for laboratory specimens and underground conditions. Laboratory experiments indicated that lava and quartzite specimens undergo measurable creep strain at stress magnitudes below the failure stress. Quartzite is however prone to more significant creep at lower stress levels. A viscoelastic analysis showed that laboratory creep results can not fully explain the behaviour observed underground. The in situ behaviour is governed mainly by the rheological behaviour of the fracture zone and the resulting time-dependent formation of new fractures. A viscoplastic discontinuity model implemented in a boundary element program is used to simulate this behaviour.

Jan 1, 1997

By Barry Voight

The law [ ] where [ ] is a measurable quantity such as strain and A and cu are empirical constants, describes the behavior of materials in terminal stages of failure under conditions of approximately constant stress and temperature. Applicable to intact rock and rock masses, the relation may be extended to conditions of variable stress and cyclic stress, and may be used to predict time to failure. Applications in rock mechanics include slope and excavation stability and various topics in ground control.

Jan 1, 1989

By K. Meyer

"The largest cost centres of the David Bell Mine concentrator are grinding, leaching and effluent treatment. Optimizing results during the 6th year of operation are listed below: o SAG mill slurry density control and operating strategies improved liner life and media consumption by 10 and 25%, respectively.o Poly-met liners clad the SAG mill except for the feed end. Lower noise levels and reduced mill downtime have resulted in an improved and more efficient workplace (time required for liner changes is 2 to 3%).o A comparative grinding media study using seven different marked media at the same time showed media wear rates range 120 to 198 mg/kWh and ball mill charge replacement frequencies of 8.3 to 13.1 per year, and that a meaningful wear test must include all media to be tested at the same time in the same mill. High quality carbon steel cylinders had the best economics.o Pre-aeration, using pure gaseous oxygen is accomplished in pipe reactors. Small bubble generators provide oxygen in the first three leach tanks. 1500 to 3000 kg of oxygen is consumed per day. Leach kinetics have improved and trends indicate increases in gold recovery of three to seven ounces per day at reduced costs. Effects of reduced cyanidation of accessory metals on the effluent treatment plant will be discussed in a separate paper."

Jan 1, 1991

(Published by courtesy of the Great Boulder Pty. Gold Mines, Limited)1. INTRODUCTIONThe faults described in the following paper are of a somewhat unusual character, and have presented a serious problem both in the development and mining of many ore-bodies on the Eastern side of the Boulder Belt. They are mentioned only briefly in the literature of the field, but were observed by, both Stillwell (1)t and Gustafson (2), and have been studied by other geologists and operators in various parts of the "Belt." Their effect on the lodes of Hamilton Section, Great Boulder Mine, was described by King (3) in 1937, and, during recent years, they have been encountered in most of the Great Boulder workings extending east of the Boulder Dyke.Although this paper is based mainly on a study of faulting and development problems on the Great Boulder mine, the general conclusions reached may be applicable to the solution of similar faulting problems in other mines throughout the field.In order to arrive at a clear understanding of the somewhat complicated relationships which the east-dipping faults bear to the lode systems it will be necessary to give a brief outline of the salient geological and structural features of the "Belt" before proceeding to describe the faults themselves.2. GENERAL GEOLOGYThe most recent descriptions of the geology of Kalgoorlie are those by Stillwell (1) and Gustafson (2) and the reader is referred to these as representing the basis of current knowledge of the geology of the field.The Boulder Belt is comprised of a central core of altered quartz dolerite which is flanked on either side by fine grained greenstones and lavas, usually referred to as the calc schists. Down the centre of the quartz dolerite mass there is an infolded bed of slate (2, p. 105) which has been intruded in lit par lit fashion by albite porphyry; the composite body of slate and porphyry so formed is known as the "Boulder Dyke." Numerous smaller porphyry dykes intrude both the quartz dolerite and the calc schist; these vary in composition and range from chloritised hornblende porphyrites to albite porphyries.The whole of the rock assemblage has been subjected tG a considerable amount of folding, faulting, shearing and fracturing, the greater part of which took place before the formation of the lodes.

Jan 1, 1941

By P. Chaddock

The Thompson Mill has a capacity of 15,000 T ID of ore and produces two grades of nickel concentrate, copper concentrate and occasionally graphite concentrate. Up to July of 1985 we processed two distinct ore types: T-1 ore, averaging 2. 8% Ni and . 2% Cu and Pipe Open Pit ore, averaging . 8% Ni associated with peridotite and serpentine. Presently our ores from T-1 mine and Thompson Open Pit are of the same type and are processed together. (Figure 1)! n October, 1982, we decided to install an Outokumpu Courier 300 on-stream analyser on an initial 6 streams. The equipment was delivered to the mill on May 2, 1983 and the first analyses were generated by June 17th. The calibrations were 35% complete by the mid-July summer shutdown and were completed in the fall. 1. 2 Analyser Slurry Feed System:The key to successful on-stream analysis is delivery of the sample streams to the analyser.Our system uses pressure-thief sampling for collection. (Figure 2) Initially we used thief samplers modelled after the Outokumpu design, with turbulence bars upstream of the collection point. These units have been replaced with simple laterals as they have worn out. Our thief samplers are mounted vertically on pump discharges and we believe sufficient turbulence exists to ensure representative sampling.

Jan 1, 1987

By O. R. Eckstrand, M. Asif, P. L. Schwann, A. J. Naldrett, R. J. F. Scoates

Paper presented at the IAGOD international symposium on mineralisation related to mafic and ultramafic rocks, held in Orleans, France, 1-3 September 1993. PGE-bearing sulphide mineralisation is contained in the lower 600 m of the Upper Central Layered Zone of the intrusion. There are four regimes with distinctive crystallisation characteristics. PGE content is low, both in bulk rock and in terms of metal in 100% sulphides, in the sulphide-rich cumulates but increases sharply in the sulphide-poor cumulates at the base of regime 1. Increases in regime 2 have no such clear relationship with incoming cumulus sulphides. PGE contents are low throughout regime 3, but regime 4 has sulphides with high PGE content, decreasing rapidly upwards, at the base of many of its cyclic units. The observations are interpreted as indicating that sulphide saturation occurred in the middle of a cyclic unit in regime 1, after which fractional segregation of sulphide caused a systematic decrease in PGE content. The resulting contrast between 40 000 ppb PGE in sulphides below the mineralised zone and less than 1 000 ppb above is of practical importance in exploration for such zones. In regime 2 the magma chamber was open to fresh influxes and PGE content was not so sensitive to sulphide saturation. In many regime 4 units influxes of new magma caused immediate sulphide saturation and PGE depletion, so that PGE-rich cumulates do not underlie PGE-bearing sulphide zones. Theoretical analysis indicates that regime 1-type behaviour is likely to characterise intrusions below the level at which plagioclase has entered as a cumulus phase and regime 4-type behaviour above

Apr 1, 1994

Overcasts are indispensible ventilation control devices to permit one airway to pass over another without mixing in room and pillar coal mining. The airflow distribution inside the overcast depends on its spe-cific configuration at which the flow enters into it. This study uses 3D numerical simulation to examine the pressure and velocity distributions inside an overcast. The effects of the overcast angle on the flow field are also examined. The flow is assumed to be turbulent and isothermal and the standard k-e model is utilized. Simulated results display flow separation and recirculation regions of different sizes occurring not only upstream and downstream from the overcast but also inside it. The results also show that the size of the recirculation regions in-creased with the increase in the angle before and after the overcast influencing significantly the flow field and pressure losses. The stream-lines are plotted to show the 3D structure of the flow field. The results from the simulation can guide overcast design commonly used in un-derground coal mines. Keywords: 3-D simulation, Overcast, Room and Pillar, Turbulence, Recirculation Region

Feb 27, 2013

By H. K. Lin

Deepwater discharges of sulfide flotation tailings have been practiced for several decades. Oxygen penetration and organic matter content in such environments are the two major factors that determine their reduction-oxidation condition and hence the rate of sulfide mineral dissolution under such discharge environments. In this study, the effects of solution redox potential and pH on the dissolution of some heavy metals from sulfide flotation tailings were examined. Zn, Pb and Cu concentrations in solution were all higher at a higher redox potential and lower pH. Dissolved Zn concentrations reached 100 ppm even at a low redox potential of 105 mV and near neutral pH after 33 days. Pb concentrations were below 2 ppm at 105 mV and near neutral pH after 165 days. At 105 mV, the Cu concentrations were below 0.2 ppm at near neutral pH after 165 days. At pH 2.3, the Fe concentrations were higher at 420 mV than at 105 mV. But at pH 6.8, the Fe concentrations were lower at 420 mV than at 105 mV, due to the oxidation of Fe2+ to Fe3+ and the subsequent precipitation of ferric hydroxide. Arsenic concentrations in solution were all less than 0.002 ppm. For Zn, Cu and Fe, no significant differences were observed between the tests in solutions prepared using seawater and those using distilled water. However, Pb concentrations reached higher levels in solutions prepared using seawater.

Jan 1, 2009

By AIME AIME

EDWIN W. KEMMERER, professor of international finance at Princeton, in a speech before a banking conference at Urbana, Ill., on Nov. 26, stated that the increase in the store of gold held by the United States is not so alarming as it appears on the surface. Excerpts from his address, as reported by The New York Times, follow: "It is significant . . . that the value of gold as expressed in its purchasing power in the United States during the last half-dozen years has been higher than at any other equivalent period since our records began 140 years ago. Our so-called 59-cent dollar today buys about 25 per cent more goods at wholesale than did our old dollar in the prosperous year 1926, and an ounce of gold will now buy considerably more than twice what it did then. This present high value of gold is one reason for its recent large production.

Jan 1, 1941

Wetting Characteristics of Mineral Surfaces - Contact Angle Measurements through Molecular Dynamics Simulations

Sep 13, 2010

A fundamental component of grade control at Tarmoola Mine has been the monitoring of assaying through quality assurance programs. The programs have concentrated on issues related to accuracy and precision of leachwell and fire assay techniques, sample preparation and grade variability within a sample. An outcome of monitoring has been identification of shortcomings in the application of assaying techniques to Tarmoola ore types that contain coarse particulate gold. Leachwell has shown satisfactory precision since introduction as the primary assay technique in October 1998 but leachwell accuracy and recovery through comparison with fire assay techniques has identified bias. A bias toward higher fire assays in grades above 2 g/t invoked an investigation into the cause of the discrepancy between leachwell and fire assay. Testwork evaluated Leachwell assay recovery by varying leach time duration, which showed that by increasing the leach time improved recovery significantly. Monitoring of leachwell against fire assay had shown that a leachwell (200 g) with a one hour leach has under-called grade when compared with fire assay by up to 18 per cent, on average, for Tarmoola ore types. Alternatively, optimisation of the leachwell assaying technique using a 200 g pulp and a two hour leach has returned, on average, higher assays than fire assay by up to eight per cent. Testwork on screen fire assays, duplicate assays and fire assay to extinction has demonstrated assay variability associated with coarse, particulate gold. The variability and distribution of gold in Tarmoola ore types has emphasised the importance of high quality sample preparation for leachwell assaying to ensure liberation of gold to leaching solutions. Assay quality assurance protocols are vital to assay data integrity and enable confidence in delineation of ore-waste boundaries and provide early recognition of problematic assaying periods and isolation of grade control practices, which depart from performance standards critical to the operation.

Jan 1, 2000

By Johnston W. H

MINING operations in the Waihi mine began, under the present proprietorship in the year 1889. Previous to this, and as far back as the year 1880, the presence of reefs in the Martha Hill had been known, and surface working on a small scale had beencarried out, without any great measure of success. The only method of treatment in use at this time was the old system of light stamps, amalgamating tables and berdans, and recovery by this method being low, little margin for profit was left, so thatthe property in these days was not considered of great importance and work on the area was only carried on spasmodically by different owners.On the transfer of the leases to the present company, calcining, dry crushing, and hot pan amalgamation was adopted as the method of treatment, and this system was continued till the more modern cyanide process was introduced and applied. As themethods of treatment and gold-saving became more efficient with the application of more modern and scientific processes, the importance and future prospects of the mine became more and more pronounced, and mining on a larger and more comprehensive scale was gradually established.As in the case of ore treatment, mining methods in the early stages were not as far adv.anced as the adopted practice of today, and numerous shafts, levels close together in vertical height.

Jan 1, 1912

By Jack Eloranta

This paper documents the relationship between material handling and processing coete compared to blasting cost. The old adage, "The cheapest crushing is done in the pit", appears accurate in this c!ase study. Comparison of the accumulated cost of: powder, selected wear materials and electricity; indicate a strong. inverse correlation with powder factor (lbe powder/long ton of rock). In this case, the increased powder cost is more than offset by electrical savings alone.

Jan 1, 1995

By Eric S. Lindquist

High groundwater levels combined with variable soil and rock profiles present challenges to the design and construction of deep excavation support systems. Increasingly, secant piling techniques are the preferred technical solution for shoring in these cases. Recent developments in drilling equipment, tooling and procedures allow cost-effective construction of deep, overlapped pile systems to tight tolerances in extremely difficult ground conditions. This paper summarizes critical design considerations for secant pile shoring systems, detailing recent advances in installation procedures and verification techniques that allow secant piles to be used for an expanding range of project conditions. Technically- and commercially-viable secant pile shoring designs require that stringent drilling tolerances be achieved in order to successfully provide combined structural support and groundwater cutoff systems, particularly for unreinforced circular secant pile shafts acting in ring compression. Modern drill rigs and tooling can advance fully cased shafts in geotechnical conditions which range from cohesionless soils below the groundwater table to hard rock, within a single drillhole. Down-hole instrumentation provides verification that tolerance requirements are met and allow secant pile shoring to be used at sites previously considered unacceptably risky. In California, secant pile shoring systems are being used for excavation depths of over 100 ft (30 m) on projects with challenging site conditions. Two case studies are presented detailing severe geotechnical conditions and excavation depths which represent the regional state of practice in secant piled excavation support. These projects provide real world examples of the design, construction and construction verification techniques that permit secant pile shoring to be used in more demanding geotechnical conditions and to greater depths than ever before.

Jan 1, 2011

By W. R. Hardwick

This Information Circular describes copper mining and concentrating methods, practices, and costs at the Silver Bell Unit of American Smelting and Refining Co., Pima County, Ariz. It is one of a series being prepared by the Bureau of Mines on mining methods, practices, and costs in various mining districts of the United States. The Silver Bell mining district, one of the older districts in Arizona, is in the western part of the Silver Bell Mountains in Pima County, about 40 miles west of Tucson and about 20 miles southwest of Red Rock, a station on the main line railroad of the Southern Pacific Co. The Oxide and the El Tiro open-pit mines of the American Smelting and Refining Co. are in the district (fig. 1. The principal ore deposits are in T. 12S., R.8E., Gila and Salt River base line, at an altitude of about 3,000 feet. The present town of Silver Bell, established in 1952, is reached by a paved highway, the Avra Valley road, that turns west from Highway 84 at Plata siding near Rillito and extends westward 25 miles to the townsite (fig. 2). The present town is about 5 miles southeast of the old town of Silver Bell that flourished from 1900 to 1930.

Jan 1, 1963