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By J. C. Draper

This subject was introduced last April in a paper presented in Birmingham, Alabama, by James E. Elkin, General Superintendent of Duquesne's Coal Department. This paper will cover some of the same data and in addition it will update and further describe our experience. Our Portal No. 3 was started in June 1965 with battery tractor-trailer face haulage. This mine is operating in the Sewickley seam in Dunkard Township, Greene County, in Southwestern Pennsylvania as shown on Fig. 1. The Warwick group of mines fronts on the Monongahela River and extends westward for some ten miles as shown on Fig. 2.

Jan 1, 1968

By Aydin Shaterpour Mamaghani, Bilgin. Nuh, Engin Dogan, Tayfun Erdogan

"Improved safety and more environmentally friendly operation are basic advantages of mechanical excavation over the drill and blast method. Raise Boring Machines (RBMs) are one of the types of these machines which have recently been increased use in Turkey. The excavated hole with RBM is more stable than other methods and has better air flow, making it ideal for ventilation shafts. This paper is a summary of the performance of a RBM in the Balya PB-Zn Mine in Turkey. Length and diameter of the ventilation shaft are 200 m and 2.4 m, respectively. The geological and geotechnical characteristics of the strata excavated in the mine are first given and later the validity of a model developed to predict the machine performance is discussed. It is believed that the results of this study and the model given to estimate the performance of the RBM will be useful guide for future applications. INTRODUCTION Raise boring technology was developed to meet the demands of the mining industry, but has also found numerous applications in tunneling or infrastructure projects for ventilation purposes. Raise boring machine uses a small diameter drill rod, around 230-350 mm to drill a pilot hole down to the required depth up to around 2000 m. Once the pilot hole has been drilled to the desired depth, a reamer of up to 9 m diameter is attached to the drill rod. The reamer is then pulled back up to the upper level, creating a round shape. Continues operation provides a faster advance rate than other methods. RBMs are suitable both in hard and soft rock. The flexibility in different angles and diameters is a great advantage compared to drill and blast excavation method. RBM creates a shaft with smooth walls which usually does not require lining. The hole is more stable than a drilled and blasted method and has better air flow, making it ideal for ventilation shafts. Raise boring is convenient method in competent rock; however several problems are faced in geological formations where geological discontinuities are dominant. Moreover, assembly and disassembly time may be long in some cases. The performance and operational parameters of the excavating machine should be predicted prior to starting the excavation operations in order to fulfill the economic requirements of the project. In the light of this fact, the main objective of this investigation is to summarize a case study on raise boring machine performance and suggest a new model based on laboratory test results. First, general description of the mine and project are provided. Then, the physical and mechanical characteristics of the samples taken from project area are discussed. In the third section, performance and operational parameters of RBM are investigated to understand the machine behavior in different rock characteristics and also to collect a useful database for further investigation. In the fourth section, a new model for performance estimation of RBM are recommended."

Jan 1, 2016

By R. M. Powers

The Florida Institute of Phosphate Research (FIPR) has funded a four-year research project to develop a computer model 'which will predict the hydrologic impacts of phosphate mining on surface and groundwater systems. The core of the model is an analytical geographic information system (SPANS) which integrates EPA's Hydrology Simulation Program Fortran (HSPF) surface water model with the McDonald-Harbaugh groundwater model (MODFLOW). The hydrologic model conducts both event and continuous annual Simulations driven by variable precipitation and temperature data. The model is currently being calibrated and will be ready for distribution to the user community in mid-1991.

Jan 1, 1990

By P. B. Moran, J. W. Rozelle, J. Forbort

"The Mount Todd gold mine is located approximately 50 km (31 miles) north of the town of Katherine in the Northern Territory (NT) of Australia just off the Stewart Highway, and was operated in the 1990s until 2000 (Fig. 1). After cessation of mining operations, meteoric precipitation and acid rock drainage (ARD) associated with surface water runoff continued to report to the retention ponds across the mine site. In 2000, the mine site entered receivership and the NT government became responsible for the care and maintenance program. In 2006, Vista Gold Corp. (as Vista Gold Australia) acquired the rights to resume exploration on the mine site without the financial responsibility of the environmental liability. In January 2007, Vista Gold assumed the care and maintenance activities on behalf of the NT government, with the environmental liability remaining with the NT government, with the goal of reinitiating mining operations. A critical path component of the planned upgrades to the existing mine/milling facilities and infrastructure was the development of a cost effective treatment and dewatering program for the approximately 11 GL (2,900 million gal) of ARD stored in the Batman Pit (RP3), which has a maximum capacity of approximately 12 GL (3,200 million gal). The Batman Pit must be dewatered in order for Vista Gold to restart mining activities. The Batman Pit Lake had reached a maximum water depth of 104 m (341 ft) when Vista Gold initiated water level monitoring. It is important to note that the ARD water in the Batman Pit has been pumped into it from 2000 to 2012 under an NT government-approved environmental control program.The Batman Pit has virtually no connection to the local ground water regime due to the intensive silicification that accompanied the mineralizing event. In October 2012, just prior to treatment, the pit lake profile was characterized by uniform acidic pH (approximately pH 3.2) and oxic conditions (dissolved oxygen greater than 7 mg/L with depth (Fig. 2). The temperature decreased by approximately 4 °C (39 °F) between the lake surface and approximately 20 m (66 ft) deep, and remained near constant with increasing depth. This temperature difference is believed to be associated with an existing shallow mine bench at approximately 20 m (66 ft) deep around the perimeter of the pit. Similarly, electrical conductivity, an estimate of total dissolved solids, decreased by about 85 microSiemens per centimeter (µS/cm) from 2,800 to 2,715 µS/cm between 10 and 20 m (33 and 66 ft)."

Apr 1, 2016

By D. H. Carlson, J. H. Johnson, C. F. Renders

Introduction Diesel-powered vehicles are used extensively in underground mines throughout North America. The bulk of the diesel vehicles found in underground mining operations are used for loading and ore haulage, as well as for transportation of personnel and supplies. Along with the advantages of using diesels underground is the disadvantage associated with diesel-tailpipe particulate-matter emissions (DPM). The concentration of DPM in the ambient air of US underground metal mines is not now regulated by the Federal Mine Safety and Health Administration (MSHA). However, recent studies have shown DPM to be mutagenic (National Institute of Occupational Safety and Health, 1988), and the American Conference of Governmental Industrial Hygienists (ACGIH) has recommended that the exposures of per¬sonnel to DPM be limited to an 8-hr time-weighted average concentration (threshold limit value or TLV) of 0.15 mg/m3 (Anon., 1995). The authors, while making measurements in a number of US underground mines that use diesel haulage equipment, found mine air DPM concentrations ranging from 0.2 to 2.36 Mg/M3 (McCawley and Cocalis, 1986; Watts et al., 1989; Cantrell et al., 1991; Haney, 1992; US Bureau of Mines, 1992; Watts, 1992; Watts et al., 1995). If the proposed DPM TLV were to be adopted as a permissible exposure limit (PEL) for US underground mines, the proposed limit of 0.15 mg/m3 PEL would be lower than any of the concentrations measured in the earlier studies and would represent more than a 15-fold reduction from the maximum 2.36 mg/m3 concentration. A 0.15 mg/m3 PEL would also represent a 4.5-fold reduction from the average 0.68 mg/m3 measured mine ambient air DPM concentration reported in this paper. Other diesel tailpipe emissions that are now regulated underground include carbon monoxide (CO), with a PEL of 50 ppm; nitrogen dioxide (NO,), with a PEL of 5 ppm; nitric oxide (NO), with a PEL of 25 ppm; and sulfur dioxide (SO,) with a PEL of 5 ppm. Because the concentrations of these gaseous pollutants and DPM are affected by the state-of-maintenance (Waytulonis,1992), it is important that a means be developed to measure emissions from engines that are now in service to determine when maintenance is needed. The current study was the result of an inquiry by mine¬maintenance personnel who had been receiving complaints about high concentrations of diesel soot (DPM) in mine headings from load-haul-dump (LHD) vehicle operators. Mine-maintenance personnel were searching for an objective test to determine if the diesel tailpipe particulate emitted was excessive. The mine was also evaluating electronically controlled, two-cycle, naturally aspirated, direct-injection diesel engines on some of their JCI (John-Clark Inc.) load-haul-dump (LHD) vehicles. These LHD vehicles were used to haul freshly blasted ore from mine headings to a feeder breaker. The feeder breaker breaks down the larger chunks and feeds the broken ore onto a conveyor. Michigan Technological University, in past studies, developed an emissions-measurement apparatus (EMA) ca¬pable of measuring diesel vehicle tailpipe pollutant concentrations (Chan et al., 1992; Chan et al., 1993; Carlson et al., 1994). At the time of the study reported here, most of the mine's LHD vehicles used a 12-cylinder, four-cycle, naturally aspirated prechamber diesel engine. The study was undertaken in cooperation with mine maintenance supervisors from late 1992 through July 1993. The objectives were to compare diesel exhaust emissions between the 6-cylinder, two-cycle, electronically controlled, direct-injected diesel engine and the 12-cylinder, four-cycle, prechamber diesel engine and to, then, use the data collected, in conjunction with mine ambient air measurements, to demonstrate the application of the "deterioration factor" (Chan et al., 1992), which is a measure of the state-of-maintenance of mine-vehicle engines that are now in service. The information would be used to identify vehicles that need maintenance to reduce emissions. The data reported here are unique in the sense that they combine underground diesel vehicle ambient and tailpipe

Jan 1, 1999

By Paul R. Smith

The theoretical requirements for the flotation d large particles are discussed. A flotation machine is described which has components which meet many of the theoretical requirements for the flotation of large particles. The machine's salient features include multiple fine-bubble attachment, froth level feeding, quiescence, shallow depth, and short flotation time. Test conditions and results are presented for the laboratory flotation of 10 mesh phosphate, 4 mesh potash, and 1/2 inch coal. In addition, test conditions and results of pilot plant flotation of minus 14 plus 35 mesh phosphate and minus 14 plus 150 mesh phosphate are also presented.

Jan 1, 1969

By R. T. Torrie, J. R. Mernagh, D. B. Chambers

l, INTRODUCTION Uranium mine and mill workers are exposed to external gamma and beta radiation fields from radioactive ore. In the past the average uranium content of ores mined in the United States and Canada ranged from about 0.1% to 1.0% U308 with pockets of much higher grade ore. Holiday (1973) reported that radiation surveys in the U.S. uranium mines found mean gamma "radiation rates ranged from 0.20 to 0.70 mrem/h. Such radiation rates cause relatively insignificant exposures." Others also concluded that the external gamma radiation fields associated with uranium mining did not result in significant worker exposures (Federal Radiation Council, 1967; Simpson, S.D. [et al], 1959) Gamma exposure levels in most modern-day Canadian uranium mines are reported to be low with average annual exposures estimated to be less than 1 rem/a (Frost, S.E. [et al], 1981). However, two developments are taking place which affect the potential significance of external radiation fields in the uranium mining environment. The first development stems from the most recent system of dose limitation developed by the ICRP which is intended to limit the workers' overall risk from exposure to ionizing radiation through the adoption of a sum rule (ICRP 26, 1977) which combines external and internal radiation exposures. In the case of exposure to radon daughters the sum rule will have the effect of reducing the annual exposure limit below the recommended limit by an amount that depends on external radiation and other sources of internal exposure such as the inhalation of ore dust (ICRP 1980). The Atomic Energy Control Board of Canada (AECB) is reviewing this subject and is expected to produce its recommendations shortly. Irrespective of the form of the sum rule eventually adopted by the AECB, it is clear that the net effect of the sum rule will be a collective reduction of the individual dose limits for individual exposure pathways. The second factor is the increasing development of high grade uranium deposits in Northern Saskatchewan. Some of these ore bodies have an average ore grade of 1% to 5% U308 or greater. Since the potential external radiation fields increase in proportion to the ore grade, it is apparent that increased effort in radiation protection planning is required in order to develop safe yet workable methods for mining and milling such ores. This paper is intended to provide information which can be of assistance in the formulation and development of a mining and milling plan. The principal focus of the paper is source identification and the design of radiation protection measures to limit external gamma radiation exposure. The exposure of workers to external beta radiation fields is also discussed. The paper is organized as follows: - Section 2 deals with source characterization. - Section 3 discusses the effects of finite source size and distance (i.e. geometry effects). - Section 4 presents selected data that are useful in evaluating shielding requirements. - Section 5 discusses the potential beta radiation fields. - Section 6 discusses practical data requirements for worker exposure scenarios. - Section 7 presents a variety of work exposure calculations. - Section 8 is a summary of this paper. 2. SOURCE CHARACTERIZATION This section develops the basic formulae for estimating the fluxes and doses from external gamma radiation. The calculation of the radiation flux due to a distributed source (i.e. a linear, area or volume source) as a function of distance assumes that any distributed source can be treated as a summation of point sources. [ ] Uranium ore contains radionuclides from both the decay chains of U-238 and U-235. In this paper the radioactive daughters are assumed to be in secular equilibrium with the uranium parent. (If natural thorium were present in the ore in significant quantities, the gamma rays originating from Th-232 would have to be added to the gamma rays from the uranium series). In all, there are over 50 separate gamma rays (as well as alpha and beta particles) emitted from the U-238 and U-235 radioactive decay series (USHEW 1970). The total

Jan 1, 1981

By David A. Summers, Clark R. Barker

INTRODUCTION The use of high pressure water as a means of excavating valuable mineral from the earth dates back many hundreds of years, but is most familiar through its use in sluicing the gold bearing deposits of California and Alaska. Interestingly, that practice, which dates back from at least 1852, is now nearly defunct with only one known surviving hydraulicking operation.(l) In that particular development water under a relatively low, natural pressure head is carried to a nozzle and from the nozzle it is thrown a distance of perhaps 20 ft to wash the gold bearing material down into a channel from whence it's carried over some mineral collector, and the water is then recirculated. Development of this practice in coal mining began in Russia in 1915, and Dr. Muchnik established the criteria for hydraulic mining during the 1930's. The first water jet mine was established underground in 1939 and has been followed by several larger mines to the present day where there are at least 11 large mines producing about 10 million tons of coal per year in the Soviet Union.(2) Water jet mining has expanded from Russia through Japan, Germany, Poland, Czekoslovakia, and mainland China. The nearest location to the United States is currently at the Kaiser Mine in Sparwood, British Columbia, where a three-man crew has mined up to 3,400 tons of coal in a 6-hr shift. There are many advantages to this particular operation. It is one of the safest mining operations for coal in Canada and has won safety awards because of this. The process is relatively cheap since the cost of the equipment is much lower than that which would be required conventionally using either longwall or shortwall mining, and transportation is achieved by fluming the coal to a central station using a gravity feed and then pumping to the surface, (3) For this

Jan 1, 1976

By LeEtta M. Shaffner, John R. Marks

INTRODUCTION In the last twelve years, underground stoping methods at the Homestake Gold Mine have evolved from open cut-and-fill with jacklegs, cribbed raises and electric slushers to ramp-based mechanized cut-and-fill (MCF) and vertical crater retreat (VCR) with dieselpowered loaders, trucks and drill jumbos. The evolution was swift. Unfortunately, ventilation practices have had trouble keeping pace. Stope ventilation is still too dependent on auxiliary fans and coolers. A single large MCF stope might contain up to eight headings, each of which requires significant ventilation resources when a diesel loader is present. Air is shortcircuited all too frequently through mined-out VCR panels. These factors and the recent rapid decent of the center of mining prompted a study to improve ventilation practices and to more accurately project future requirements. BACKGROUND In 1992, the underground portion of the Homestake Gold Mine produced 8336 kg of gold from 1407 ktons of ore (268,000 oz from 1,551,000 st). The ore is hosted in low to medium grade meta-sediments. Stoping took place from 420 to 2347 m below the surface (1400 to 7700-ft levels). The weighted center of mining was 1662 m deep (5450-ft level) in 1990 and is projected to be 1890 m (6200-ft) in 1993. The deepest level is 2440 m (8000-ft) where the virgin rock temperature (VRT) is 56.1°C (133F). The mine is ventilated by 504 m3/s (1,069,000 cfm) measured at mid-exhaust-circuit density. The air-conditioning system includes an 8.1 MWR (2300 ton) controlled recirculation plant, a 2.0 MWR (580 ton) chilled water plant, a 1.0 MWR (290 ton) exploration drift refrigeration plant, 28 spot-coolers totaling 3.4 MWR (960 tons) and 35 spray coolers totaling 1.5 MWR (420 tons). The mine employs 117 diesel units with a total nameplate rating of 7961 kW (10,672 hp). These units include twenty-four 1.5m3 (2-yd) loaders, twenty-six 2.7m3 (3.5-yd) loaders, fourteen 3.8m3 (5-yd) and two 7.6m3 (10-yd) trucks, and assorted utility vehicles and drill jumbos. THE STUDY In April 1991, the University of Nevada-Reno (Mackay School of Mines) and the South Dakota School of Mines and Technology cooperated with Homestake on an MCF study. Mackay instrumented a stope with thermocouples, air velocity meters and hygrometers (Duckworth, 1992). South Dakota Tech conducted a finite-elements computer analysis of a back-filled MCF stope with/without light-weight shotcrete insulation on the sidewalls and back (Chellam, 1992). Homestake conducted an empirical analysis of deep-level MCF stoping. This paper describes the Homestake study. Twenty-three of the forty MCF stopes deeper than 1800 m were surveyed at least once during the last quarter of 1992. The stopes not included were being cable-bolted or back-filled. Figure 1 shows the complexity of one of the ramp-based stoping areas included in the survey. This particular stope has six separate production headings. A baseline wallrock heat load was derived for each stope or heading from psychrometric calculations. Broken ore, waste rock fill and fissure water were noted when present and the effects included in the baseline heat load. Other heat sources often mentioned in the literature such as metabolic heat from workers, heat from explosives and small electric loads were neglected. Fan heat was considered part of the ramp & crosscut heat load and thus not included in the study. Diesel heat was addressed separately. RESULTS Survey results, plotted as the heat flux against VRT, are shown in Figure 2. The equation for the regression line is: W/m2 = 2.1236*VRT - 75.405 The 0.31 correlation coefficient is poor which implies that the results should be used cautiously. Previous experience strongly suggests that differences in productivity are most likely responsible for the scatter in data points. A rapidly advancing stope will have a

Jan 1, 1993

By Samer Sadek, Phoebe Cheng, Changmo Kwon, Myat Ohn

BART Facilities Standards (BFS) Structural Criteria is a living document that governs planning, design, and construction of all BART projects. Since its development, the document undergoes incremental changes and updates to reflect policy changes, changes in local, State, or national design codes and guidelines, advancements in engineering and technology, or based on the performance of past built infrastructure. In 2020, the BFS-Structural Criteria underwent a comprehensive update and was released with BFS version 3.2. This paper summarizes the updates to the Tunnels and Underground Structures Section of the Structural Criteria, and provides insights to the basis and logic of the requirements.

Dec 1, 2022

By J. E. Carr

A bewildering number of continuous thickeners and clarifiers are used in mineral industries and other heavy industries. General types are: 1. Conventional Thickeners (Fig. 1). 2. Thickeners with Flocculating Feedwells (Fig. 2). 3. Solids Contactors (Fig. 2). 4. Tray Thickeners (Fig. 3). 5. Lamellas (Fig. 4). 6. Conventional Thickeners with Lamellas 7. High Flow, Fluid Bed Thickeners (Fig. 5). These units must be adapted to varied applications. They must provide operating flexibility to meet process needs. Such factors multiply the decisions which must be made prior to the selection of a thickener. The seven types of thickeners listed above, with their variations, share design concepts which limit their application and which can be used to evaluate their performance. Some of these are listed: 1. Feed velocities must be limited and/or controlled. 2. Natural or reagent flocculation is necessary. 3. Horizontal settling area is the basis for capacity. 4. Overflow velocities must be limited and/or controlled. 5. A controlled underflow is necessary. Applying these needs to the basic process concepts is thickener selection. The basic function of a thickener in the minerals industry is the separation of a thickened sludge containing fine solids from water. To this primary separation function are frequently added others to give a list as follows:

Jan 1, 1977

By Frank F. Rahne

Butler Taconite, operated by the Hanna Mining Company, is an Iron ore pelletizing operation that is located on the Western end of the Mesabi Iron Range in Northeastern Minnesota. The operation mines low-grade taconite ore and through primary crushing, two stage grinding, magnetic separation, agglomerating and indurating, produces 2.6 million annual tons of high quality, 66% iron pellets. Mining and stripping is done with 13 - M100 trucks and 5 - 12 cu. yd. shovels. The operating processes in the concentrator and pellet plant are primarily con¬trolled through a centralized control room and monitored by a PDP l140 computer. During 1969 and 1970, a re-organization of the Company took place. Through this re-organization, each mine unit was analyzed and studied to determine the proper staffing of the internal departments. At this time, it was decided to continue with further detailed studies of the Maintenance Department with the intent to reduce maintenance costs. Maintenance costs at the time of the re-organization were 38.6% of the total mine operating cost. The preliminary study indicated that the Maintenance Department had a system of maintenance, as a matter of fact, it had a separate system in each sub-maintenance division and in some of the sub-divisions each foreman had a system of his own. Therefore, from the data collected, it was decided to develop one system of maintenance that could be utilized throughout the entire operation. To determine the type of system suitable to all phases of the operation, detailed time and method studies were conducted. All types of maintenance tasks performed at the operation were studied. The data collected was analyzed and the recommended improved methods were reviewed with the Maintenance Superintendent and his staff. The new improved methods that were accepted as feasible, were introduced into the system immediately.

Jan 1, 1975

By B. H. Gardner, S. Serata, D. Dickie

Since 1958, the Sifto Salt Division of Domtar, Inc., has operated the Goderich (Ontario) salt mine in Michigan basin bedded halite, with shafts located on Lake Huron's eastern bank. Initial mining utilized a conventional room and pi1lar pattern. However, roof deterioration developed. Incidence of progressive upward failure of saltback and over1yings ha1ey-do1omite layers intensified with entry aging. Studies were therefore initiated on new mining patterns utilizing the Serata Stress Control Method. Pursuant thereto, three test areas, indicated in the mine plan of Fig. 1, were designed and excavated. Full conversion to Stress Control mining was effected in January, 1983. This paper focusses on the instrumentation work in the third test area. Test Area 3 was designed to validate the excavation geometry of the new stress-controlled main entry system. The testing of that design was accomplished, in part, by comprehensive in situ stress, property, and deformation measurements on the performance of the model main entry section of Test Area 3. The paper describes: (1) the theoretical background of the SPDR/Stress Control Method; (2) specific problems and objectives addressed by the 1983 program of in situ measurements, (3) measurement results; and (4) conclusions drawn therefrom.

Jan 1, 1986

By L. J. Blatnica

Throughout the 1980's the tem Mine-Wide Monitoring System has had various meanings for manufacturers of these systems and underground mine operators alike. System suppliers had visions of grandeur in the early 80's as they envisioned each mine, that has more than a single section, installing a computer based monitoring system to provide surveillance, control and documentation of the entire mining operation, including everything from a single heat detector to the weekly payroll. In a few cases this may have become reality but the truth is these systems have quietly become a steady but unspectacular workbase for the modern mining industry of the 1990's. As it exists today the total market for underground monitoring systems may not exceed $10 million annually. Suppliers have lowered their expectations of the dollar value of this particular market. Some were entering a new market because of experience with computer based data acquisition systems. Others had been active in the mining industry for years with more traditional equipment (communication systems, methane monitors, safety equipment, etc.) and decided to become involved with Mine-Wide Monitoring Systems.

Jan 1, 1990

By Johan K. Rodenburg

A major bauxite belt was discovered and evaluated from 1969 to 1974 by P. T. Alcoa Minerals of Indonesia in West Kalimantan, Indonesia. The proven reserves of beneficiated bauxite are 800 million tons averaging 38.6% AA1203 and 3.0% RSiO2. Inferred reserves of washed bauxite are in excess of 2 billion tons. The bauxite occurs as capping of low lying hills which were formed during the post Paleogene peneplaination stage. The overburden varies from 0.6 to 3.8 meters and the nodular bauxite horizon from 2.2 to 5.2 meters. The deposits were formed by the in situ weathering of predominantly acidic and basic intrusive rocks.

Jan 1, 1984

By AKIHIRO YOSHIMURA, HIDEKAZU KATO, KOTA TAKATORI, YASUNARI MATSUNO

Chalcopyrite (CuFeS2) is the most refractory copper mineral when treated in conventional sulfate media leaching systems. This study is the first to examine the use of a dimethyl sulfoxide (CH3SOCH3), or DMSO, solution containing copper chloride (CuCl2) in the leaching of chalcopyrite. The DMSO solution containing CuCl2 could selectively dissolve the valuable metals of copper (Cu) and gold (Au) from chalcopyrite while leaving minerals with little economic value, such as pyrite (FeS2), and the penalty element arsenic (As) in the residues. Leaching chalcopyrite using a DMSO solution that contains CuCl2 offers a number of advantages, including eco-friendliness and easy operation.

Semiautogenous (SAG) ball mills are being increasingly used today for the coarse grinding of a range of ores, including gold, copper and iron. Larger mines are demanding mills with larger throughputs. This has led to some remarkable developments in mill diameters and installed powers. Conventional design principles - mills with trunnion bearings - have reached their limitations. This article describes one SAG mill with a slide-shoe arrangement and Krupp Polysius' Combiflex twin drive. In August 1997, the world's largest SAG mill with a slide-shoe arrangement was commissioned in Queensland, Australia, operating as a wet mill. The inside diameter is 10.4 x 5.1 m (34 x 17 ft) effective grinding length. The 1.36-ktlh (1,500-stph) SAG mill is driven by two Combiflex drives (twin drives with 2 x 5,500-MW or 7,375-hp installed motor power) in conjunction with a girth gear (outside diameter 11,890 mm or 468 in.). Commissioning of the SAG mill was completed with no notable technical difficulties. The plant runs continuously except for planned maintenance stoppages and process-specific reasons. As this article is written, the machine has been in operation for about 8,000 hours without major problems.

Jan 1, 1998

By Cary Hirner, Sarah Lothman, Chris Mueller, Savita Schlesinger, Mike Hanna, Ian Fitz-James

Loudoun Water is currently implementing a quarry to reservoir conversion that will improve water supply reliability in periods of drought and flexibility in water supply operations when turbidity in its Potomac River supplies are high. The Milestone Reservoir Project consists of an inter-tie to the existing Potomac River Pipeline; new pipeline and tunnel; and a pumped intake and outlet works. Planned construction involves significant risks due to underground conditions which includes hard, igneous rock and high groundwater at the creek crossing. In addition, the site is constrained by Goose Creek, an adjacent residence, active mining in a quarry next to the new reservoir, public use of adjacent park space, and the Washington and Old Dominion Trail. Working with these constraints will require close coordination with active mining operations and public and private property owners. Safety of personnel involved with construction also represents a concern with quarry high walls over 30 meters in height near the planned construction. In consideration of key projects risks, Loudoun Water determined that an innovative CMAR delivery approach provided the best opportunity to achieve its goals in quality, innovation, front-end cost and schedule certainty, and operational flexibility and reliability. What matters most to Loudoun Water is the value in having a builder participate in the design process where key decisions can be made based on consideration of risks, cost, schedule, quality, and operability and maintainability of the system. While a progressive design-build delivery model may have produced similar outcomes, Loudoun Water had already invested in substantial development of the design for the work, and ultimately, CMAR provided Loudoun Water the opportunity to retain the partnership with the design team while gaining valuable input from a contractor. Among the challenges with CMAR in the Commonwealth of Virginia is a substantial limitation on self-performance corresponding to 10 percent of the construction value of the work. A significant industry outreach program was initiated by Loudoun Water to encourage participation with incentives in contract terms that allowed for the CMAR Contractor to make a reasonable profit. This paper reviews the CMAR procurement process and discusses contract incentives that were adopted to promote the work within the constraints of contract law in Virginia.

Jun 13, 2021

By P. W. Hynes

Tg Soda Ash has operated a shortwall mining system in the Green River, WY trona district since August 1982. The system has proven to be highly productive. In an effort to reduce the amount of relatively expensive development mining and improve mine planning flexibility, a method to mine the return air tailgate during the shortwall retreat was tested beginning in April 1987. As of June 1988, four shortwall panels had been completed using the advancing tailgate, producing more than 739 kt (815,000 st). This paper discusses the effects of the advancing tailgate on shortwall production, development mining requirements, costs, strata mechanics, and ventilation. Shortwall mining is a hybrid extraction method combining the self-advancing hydraulic roof supports from longwall mining and continuous mining machines (Fig. 1). The technique has been only marginally successful at the majority of the nearly 20 installations worldwide. At present, the face at Tg Soda Ash Inc. is the only shortwall operating in the United States and the only noncoal shortwall in existence. Despite the method's record of limited success, the staff at Tg Soda Ash determined that the conditions in its trona mine were well suited for shortwall mining. The commitment to install a shortwall was made in late 1981 and the first panel commenced in August 1982. Between that time and June 1988, about 3.33 Mt (3.67 million st) of trona have been mined with the system from 15 panels. Annual average productivities have ranged from 818 t/shift (902 st per shift) in 1982 to 1.4 kt/t shift (1591 st per shift) in 1987. Overall, the system has produced an average of 1.3 kt/ shift (1450 st per shift). In April 1987, a method of mining the return air tailgate during shortwall retreat was tested (Fig. 2). The objective was to determine if a return airway could be established and maintained by mining a series of rooms and pillars about 7.6 m (25 ft) past the face support line. The method would allow a barrier pillar to be left between shortwall panels without requiring the mining of a separate return air tailgate.

Jan 1, 1990

By J. Burrows

The Chamber of Mines of South Africa is an association of major mining corporations, gold mines, coal mines, diamond mines and other metalliferous mines. It looks after those interests of its members which can best be handled on a co-operative basis. The Chamber's activities include the engagement of labour; the negotiations of all labour matters, including minimum wage rates, with recognised trade unions and associations; the carrying out of advanced technological research; representing the industry in dealings with the government and the administration of miners' training colleges and rescue training stations. Subsidiary companies of the Chamber handle matters such as the refining of gold and other metals; the processing and marketing of uranium and the marketing of Krugerrands. The thirty-two gold mines members of the Chamber of Mines of South Africa produce approximately 694 tons of gold per annum. To achieve this they employ 430000 persons and break about 8400 kilotons of rock per month. The working revenue is approximately 4 225 million US dollars per annum and the annual working profit is approximately 2 050 million US dollars. The electric power consumed by the gold mining industry is 1650 MW, about 20% of the total power produced in South Africa, and the annual costs of this power is approximately 200 million US dollars per annum. It is estimated that about 25% of this power is used for ventilation and refrigeration appliances. Thus the annual power costs for ventilation and refrigeration is approximately 50 million US dollars. The gold bearing reef is a narrow tabular deposit situated at varying depths throughout the gold fields. The weighted mean rock breaking depths and, virgin rock temperatures range from 1 700 - 9 000 ft (520 - 2 740 m) and 68 - 113°F (20 - 45°C) respectively. The maximum virgin rock temperature at which mining currently takes place is 130°F (54.5 C). Obviously at such depths and temperatures and with such a large labour force the provision of an acceptable environment is critical and to achieve this 59 000 000 cfm (28 000 m3/s) of air is downcasted during the main shift and 122500 tons (430 MW) of refrigeration are provided. Legislation in South Africa requires that on metalliferous mines when more than 1000 persons are employed underground during any one shift the manager shall appoint one or more ventilation officers. The legal responsibilities of the ventilation officers are concerned mainly with the measurement and control of dust. In practice, this is not the main part of their work and ventilation personnel perform many other duties such as measuring and controlling the distribution of air, checking the performance of fans and cooling plants, planning air and refrigeration requirements and, more recently, measuring and controlling noise and illumination levels. Ventilation personnel are specialists who provide management with advice on all matters relating to the mine environment. At the present time there are 506 persons engaged as ventilation personnel in gold mines and a further 1744 persons are employed in gold mine ventilation departments in a non-technical capacity, for example as construction workers, filter cleaners, fire patrollers, etc.

Jan 1, 1980