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By S. Veeramasuneni

The interaction of soluble salt particles in their saturated brines is of interest from both a fundamental and practical perspective. The state of dispersion/aggregation of alkali halide particles as determined by optical microscopy can now be described based on non-equilibrium electrokinetic measurements for alkali halides. These results indicate that the sign of the surface charge has a significant effect on the extent of particle dispersion/aggregation. Atomic force microscopy (AFM) is now being used to directly measure the interaction forces responsible for the stability of these alkali halide aggregates observed in optical microscopy studies. These experimental results can be explained based on the interfacial polar interaction theory. In this regard the nature of such particle interaction forces at high ionic strengths are examined to improve our understanding of these systems.

Jan 1, 1997

By R. L. Murray-Smith, M. E. A. Robertson, M. J. Prosser

The AZTEC is an automated instrument, for the non-destructive analysis of gold ore samples, which uses energy dispersive X-ray fluorescence (XRF), This is achieved by X­ray tube excitation of gold K X-rays. The first production instrument was commissioned at Vaal Reefs Chemical Laboratory (VRCL) in November 1987, followed by two second generation instruments in 1989. AZTECs have also been installed at Western Deep Levels and at Anglo American Research Laboratories (AARL). A further unit is being assembled at Ashanti mine, Ghana. At VRCL the AZTECs are used primarily for the evaluation of mine (stope) samples. Each of the production instruments can routinely analyse 650-700 samples per day, with the aid of a robotic sample transfer system. Various techniques have been evaluated for increasing sample throughput. For example, higher value samples can be counted for a shorter time to obtain the necessary relative precision.

Jan 1, 1991

By Steve O. Green

Introduction The US coal mining industry is in the midst of a serious economic crisis. To overcome increased operating costs, operators must take advantage of technological advancements in mining systems and equipment. A relatively new concept available to underground operators is that of extending the cutting depth for continuous mining machines (continuous miners). This system is referred to as deep-cut mining, and may be described as mining cuts to depths greater than the standard 6 m (20 ft). Productivity is increased through the reduction in continuous miner tram time since the machine stays longer in each working face. To successfully implement a deep-cut mining system, three problems must be overcome: Exposing personnel to unsupported roof, methane removal, and controlling respirable dust. Remote control cutting The solution to the unsupported roof problem is remote control. Recent advances in remote operation, especially in radio remote controls, have resulted in what is generally accepted as a safe and reliable tool. By installing remote control on the continuous miner, the operator and helper can now be stationed in a supported area. This allows for improved visibility while cutting coal. With this in mind, the limiting factor now becomes the position of the haulage equipment operator. North River No. 1 uses Joy 21SC shuttle cars and Jeffrey 4110 diesel ramcars. These cars are pulled against the bumper of the continuous miner. Thirteen meters (43 ft) can be cut with the shuttle car and 14 m (45 ft) with the Ramcar before the operator of the car is inby permanent supports. By loading away from the bumper, about 1.5 m (5 ft) can be added, bringing total cut depths up to 14.5 and 15 m (48 and 50 ft), respectively. Methane removal The second problem involves methane removal, and the line curtain-to-face distance. In other words, how to maintain the line curtain to allow sufficient ventilation to remove methane and not expose workers to unsupported roof. The solution, in most cases, is to use a blowing-type ventilation system. Unlike the exhausting system that, to be effective, must be installed within 3 m (10 ft) of the working face, the blowing system can be effective at greater distances from the face. Numerous studies have documented the advantages of a blowing-type ventilation system over an exhausting system. Furthermore, a blowing system can be significantly improved by in¬creasing the velocity of the air delivered without increasing the volume of air. With this greater ventilation efficiency, it is possible to secure ventilation plan approval from the Mine Safety and Health Administration (MSHA) for a greater curtain set-back distance. This eliminates the need for personnel to travel inby permanent supports to extend the face brattice. The distance from the face MSHA will approve, however, is on a mine-by-mine basis. MSHA will evaluate the particular mine's face methane liberation rates, along with the specific proposals regarding face ventilation volumes and velocities. Dust control The third problem is controlling respirable dust. This is actually a result of the solution to the second problem. In the blowing system, while mining on the slab cut, both the continuous miner and haulage units are located on the dusty side of the line curtain (Fig. 1). In the early 1970s, underground operators virtually abandoned this system, in order to comply with the 2 mg/m3 (0.87 gr per cu ft) federal dust standards. The blowing face curtain did offer increased face ventilation efficiency. With that, greater curtain-to-face distances were also

Jan 8, 1987

By Thomas M. Larimer

Prior to April, 1961, the coal charge to Clairtonts coke oven batteries was 3" x 0" mine waslied coal having approximately 6.0% to 7.0% moisture. Even at this size and moisture content, the arching, rat-holing and hang-up was excessive and the effective live storage of coal was only 60% to 70% of the rated capacity of the coal bunkers, depending upon the moisture of the coal. In addition, many production delays were encountered in the time and methods required to fill the larry cars. The need for an improved metallurgical coke led to the introduction of pulverized coal to a portion of the plant in 196 1. When the full potential of the pulverizing equipment was realized, we were able to serve seven batteries in our second unit of batteries with pulverized coal through two coal bunkers. The discharge hoppers of these bunkers were equipped with air vibrators to assist the flow of pulverized coal into the larry cars. The vibrators were helpful in this respect but required considerable maintenance and caused damage to the bin plates.

Jan 1, 1968

Conventional wisdom in the mineral processing industry leads one to believe that pressure oxidation of refractory carbonaceous (preg-robbing) ores is not a practical approach to process development. Initial acid pressure oxidation testing on carbonaceous sulfide ores at the Twin Creeks Mine supported this conventional wisdom. After testing other refractory process options, including air and oxygen enriched roasting, Santa Fe Pacific Gold decided to take a second look at pressure oxidation, using conditions outside the normal range of conventional practice. A combination of very fine grinding and high operating temperature and pressure resulted in significant increases in gold extraction. This paper details the effect of operating parameters developed for Twin Creeks on gold recovery, sulfur oxidation, carbon oxidation/passivation and preg-robb through the laboratory and pilot plant investigations.

Jan 1, 1995

By John W. Leonard

The typical tunnel boring machine in use on today's tunnel project uses 15'2-inch diameter cutters at a nominal load of 45 kips per cutter. The cutterhead rotates at a speed of 120 over D revolutions per minute where D is the cutterhead diameter in feet. The typical machine operates about 50 percent of the time. Delivery of a new machine takes 12 to 15 months. The TBM is a specialized tool and not applicable to all jobs. There is a range of job conditions that limit the economic use of a TBM. The machine should provide a thrust per cutter at least twice the compressive strength of rock in psi and up to three times the rock strength is required if in "granitic rocks." The rock strength in psi must be at least three times the cover above the tunnel in feet or the machine may stick. For example, 2,500 feet of cover requires at least 7,500 psi rock to avoid problems. When the RQD's over a significant distance of the project are less than 50, the TBM becomes support-bound. When a large percentage of the time is spent waiting while the ground is supported, the TBM cannot perform well (See Figure 1). This problem increased inversely with tunnel diameter. We are finding that increasing the RPM's does not significantly increase progress. The increased RPM results in a substantial decrease in cutter penetration. The explanation for this decrease appears to be vibration, cutterhead drag, or lack of stiffness in the cutterhead (See Figure 2).

Jan 1, 1981

By M. Tomono

Harima Works is operating as the zinc and lead simultaneous smelter utilizing the imperial Smelting Process. The smelter commenced its operation in 1966 with designed capacities of 3500 t/M of zinc and 1500 t/M of lead. After that, owing to the results of progressive developments and improvements in operating efficiency and equipments, it has doubled its production figure. This report reviews our recent developments carried out from 1983 to 1985. Firstly, the furnace bottom design was modified for the purpose of reducing carbon consumption ratio. As a result, shaft operating condition has been satisfactory and carbon consumption ratio was improved by 2%, in addition, furnace zinc production was increased by about 8%. Secondly, the oxygen enriched operation at sintering machine was tested from July 1984 to June 1985, in order to increase the sinter productivity, and to solve the capacity imbalance between sinter plant and ISF. As the test result, it was found out that sulfur elimination rate was increased by approx 10% with the oxygen enrichment of 500 Nm3/hr.

Jan 1, 1986

By Mark L. Bricker

The Jim Bridger Coal Mine, operated by Bridger Coal Co., is located in Sweetwater County, WY about 56 km (35 miles) northeast of Rock Springs. Bridger Coal is a joint ven¬ture between Nerco Coal Corp, the operator, and Idaho Energy Resources Co. The permit boundary encompasses more than 81 km'- (20,000 acres) on the western side of the Continental Divide. Since the mine began opera¬tions in 1974, it has produced about 82 Mt (90 million st) of coal while moving about 344 hm3 (450 million cu yd) of virgin earth. Annual production averages 4.5 Mt to 6.4 Mt (5 million to 7 million st) of run-of-mine coal. Coal is consumed on-site by the Jim Bridger Power Plant, a 2000-MW generating station. Electricity is transmitted to the Pacific Northwest power grid that includes Oregon, Idaho and northern California. Geology The mine is located on the northeast flank of the Rock Springs uplift. This is a broad, asymmetrical anticlinal feature about 97 km (60 miles) long and 56 km (35 miles) wide with a north-trending axis. This uplift separates the Green River Basin to the west and the Great Divide and Washakie Basins to the east. Precambrian rocks on the apex of the uplift are estimated to be 5200 m (17,000 ft) above the Precambrian rocks in the Green River and Washakie Basins. Normal faults with generally less than 30-m (100-ft) vertical displacement cut through the uplift. The dip of strata on the flanks is generally between 3° and 15°.

Jan 1, 1991

By Richard M. Winar

In September of 1988, the United States Environmental Protection Agency (US EPA) published final regulations that effect owners and operators of underground storage tanks (USTs). These regulation include stringent technical requirements for owners and operators of USTs containing petroleum and certain hazardous substances. The technical requirements effect both new (those installed after December 1988) and existing USTs and include notification; performance and operating standards; release detection, reporting, investigation; release response and corrective action; and closure requirements for out-of-service USTs. The USEPA also published final regulations regarding financial responsibility requirements for owners or operators of petroleum USTs in October of 1988. These regulations require owners or operators of petroleum USTs to demonstrate financial responsibility for the costs of corrective action and compensating third parties for bodily injury or property damage caused by releases from petroleum USTs. Like thousands of other industries, the mining industry must become thoroughly familiar with these new regulations in order to properly manage their UST systems, ensure compliance, and understand other ramifications imposed by these regulations.

Jan 1, 1990

By R. Hogg, L. G. Austin, R. C. Klimpel

Slurry densities and holdup were measured for open circuit grinding tests in a laboratory overflow ball mill of 0.3 m diam by 0.6 m long, by emptying and drying after stoppage of the mill. Contrary to some other results in the literature, there was negligible variation of slurry level in the mill as a function of flow rate over a range of increase of flow rate of about 5 to I. However there was a gradual increase of solid holdup at higher flow rates because the slurry density in the mill increased. Equations are given for holdup and slurry density in the mill as a function of over flow diam, ball load, and feed slurry density. The average level of slurry necessary to give over flow was found to be much higher than that necessary to fill the mill to the overflow level at rest. The total level of balls and slurry was only increased slightly by ball loads higher than the overflow level. The overflow diam appeared to be the major controlling factor in determining the mill holdup.

Jan 1, 1990

By Dale W. Norris

United States Steel's Cumberland Coal Operations is located in Greene County in the southwest corner of Pennsylvania. The mine and related facilities have been in operation since September of 1977. Production from this mine, U. S. Steel's first commercial coal mining venture, will eventually reach a level of 2.72 million mt of clean coal per year and is totally committed to electric power generation. Cumberland is a slope mine in the Pittsburgh seam and is mined by continuous mining machines and longwall. The coal from the mine is trans- ported by belt to the surface, taken through a rotary breaker for initial preparation, and stored in raw coal silos. It is then processed through the preparation plant and taken to a clean coal silo where it is held for train loading. Cumberland Mine operates its am 28.16 km railroad to transport the coal to the Monongahela River. The unit train consists of thirty-one 90 mt cars. The coal, once loaded on barges, is then shipped to the customer. Cumberland Mine is one of the most modern coal preparation and unit train loading systems in the nation. The preparation plant is de- signed to receive coal from the silos at a rate of 720 at per hour. The +6.35 mm material is screened out and cleaned in a Daniels heavy media vessel with two 3.35 m weirs. Minus 6.35 mm material is fed to sixteen #88 Deister concentrating tables. The cleaned coal from this circuit is dewatered, and all effluents from the mechanical dewatering report to the classifying Cyclone sump for classification through ten 35.6 cm cyclones. The cyclones overflow material (-150 mesh) is processed through two banks of three 8.5 m3 Denver D.R. flotation machines. The cyclone underflow pro- duct (28-1 0 mesh) is processed in two banks of 5 six 2.83 m Denver Sub A machines. The clean flotation product is then dewatered in Bird screen bowl centrifuges, blended with the other clean coal size fractions, and the total plant product further dewatered in a thermal dryer.

Jan 1, 1980

By R. Kramer, A. Habibi

A ventilation survey was conducted in an underground longwall trona mine. The mine is relatively shallow and categorized as a gassy mine. The ventilation system includes three intake and six exhaust shafts, and three surface axial fans ventilate the mine in a blowing configuration. Airflow quantity, pressure losses and air psychrometric characteristics were quantified, and the influence of moving cages and skips in the shaft examined. The benefits of utilizing high-accuracy digital pressure transducers and advanced psychrometers were seen when undertaking a “leapfrogging” survey to gather data needed to build a computer ventilation model. Two underground booster fans with VFDs are idle but available at the mine. The ventilation model was used to determine the optimal booster fan locations to decrease operating cost by reducing the load carried by the main fans. A leakage study was conducted to evaluate the effects on the system of additional pressure from a booster fan. The study finally prepares a future ventilation model for the next 15 years of mine operation and investigates the effect of proposed scenarios.

Jan 1, 2014

By Jianliang Li, Xi Peng, Guorong Lu

This paper describes a wind tunnel modelling study on airflow and pollutant concentration in a large open pit mine located in southern China. A 1:1000 scale model of the surface mine was fabricated and placed in a wind tunnel. Airflow velocity in the test section of the wind tunnel can be continuously adjusted from 0.3 m/s to 6.0 m/s. In this study, both the present pit condition and the projected final pit condition were modelled. Airflow direction and velocity profiles were obtained. The impact of wind direction and velocity on airflow was also studied. The pollutant concentration distribution in the open pit was simulated using smoke, and a distribution profile was obtained. Respirable dust concentrations at the mine site were measured, then its concentrations at the projected final pit condition were estimated using the obtained pollutant concentration distribution profile. Under normal meteorological conditions, it is expected that the respirable dust concentration for the final pit should not exceed the government standard. Key Words: wind tunnel modelling, open pit, surface mining, scale model, mine ventilation, airflow profile, point source.

Jan 1, 1993

By Ying Lu, Jeffrey C. Johnson, Chen-Luh Lin, Jan D. Miller, Sindhoora Puvvada

The amount of energy used to break rock is of interest in mine-to-mill operations to minimize energy consumption during blasting at the mine site and comminution at the processing plant. In this research, a direct energy measurement was made using high-velocity impact to cause breakage of long granite cores, and to study the energy dissipation along the entire bar length. The results show the formation of three zones of damage — fragmentation, fracture and elastic — that have been characterized in 3D using X-ray tomography. The fragmentation zone of the granite bars was studied in detail to describe the particle size distribution and the particle shape characteristics. The experimental energy dissipation in the fragmentation zone was found to be in close agreement with the expected energy calculated from the Bond work index (BWI) equation [1] when the anisotropic granite particles in the fragmentation zone are defined by length rather than by sieve size. The results increase our understanding of rock breakage and energy consumption and may find future applications in mine-to-mill operations.

By A. D. Randolph

Comminution represents one of the major costs of mineral recovery from low-grade ores. Efficient mineral recovery in a typical low-grade ore grinding-flo¬tation process depends on grinding the ore small enough to liberate mineralization without producing excessive slimes, which would result in higher power and flotation reagent costs. Size criteria for efficient grinding should be stated in terms of the distribution of mineralization sizes, but as the mineralization-size distribution is difficult to determine, less desirable criteria of absolute particle-size distribution (PSD) are used in practice. Thus, there is considerable incentive to be able to analyze and predict the PSD performance of grinding circuits for control and optimization of the process. Any realistic model of a grinding circuit must account for recycle of oversize material; such oversize recycle adds to the total mill load and limits total production rate. Simulation of such closed-circuit operation to study grind factors affecting total throughput can be done in a straightforward fashion provided recycle calculations can be added to an accurate input-output model of PSD in the grinding mill itself. Such a total simulation of mill performance is of obvious practical value in assessing the practicality and benefits of various over-all mill control schemes and/or modes of operation. For example, considerable interest has been shown in the development of continuous on-line particle-size sensors and such instruments should be an operating reality in a few years. A total closed-circuit PSD model of this type would be invaluable in the utilization of, and in predicting the pay-out from, such an on-line particle-size sensor. The purpose of this paper is to present a new approach to the development of an input-output PSD model of a grinding mill. The model is continuous in the par¬ticle size coordinate, L, and is based on the principle of a particle population balance along the particle size coordinate. Most previous studies of the grinding unit operation attempt to relate power

Jan 1, 1971

By Donald P. Richards, William M. Greenslade, Daniel E. Albers

INTRODUCTION The application of oil field technology to large diameter hole drilling is beginning to offer a cost effective alternative to conventional shaft sinking methods. The drilling of ventilation and small production shafts are, of course, not new. The major constraint to the drilling of large diameter production shafts has been the inability to drill holes large enough and straight enough to install and operate relatively high volume hoisting equipment. This paper discusses recent advancements in drilling technology as applied specifically to large diameter holes for use in the mining industry. The paper emphasizes the geotechnical and hydrologic considerations in assessing site conditions and their application to drilled shaft technology. The authors worked together on a recent feasibility study for a 6 m (20 ft.) diameter 975 m (3,200 ft.) drilled shaft in New Mexico. Their individual experience in drilling and geotechnical investigations related to shaft sinking is world wide. DRILLING TECHNOLOGY Large diameter hole drilling has been a natural evolvement of oil well rotary drilling techniques. Most problems that were encountered in the rotary drilling of large diameter holes have been

Jan 1, 1981

By S. Makni, F. Wang, A. Rosa, B. Klein, C. Saud, R. McIvor, A. Kumar, C. Gagnon, D. Gong

In response to the Natural Resources Canada “Crush It Challenge”, Corem partnered with the University of British Columbia and led a project to develop and demonstrate the use of high-pressure grinding to replace the workhorse of the industry, wet ball milling, with high-pressure grinding. Two plant case studies, a small tonnage gold recovery operation, and a large tonnage copper recovery operation, were carried out. Plant circuit performances were compared to pilot plants operated on the same circuit feed set up to represent the novel equipment and flowsheet that was developed to produce separation circuit feed. Numerous obstacles to industrial implementation of highpressure grinding in this role which were recognized at the outset of the work were successfully addressed. In both cases, comminution equipment energy savings alone, and final stage total circuit energy savings, exceeded 60% and 50%, respectively. Downstream mineral separation performance was the same or better.

Mar 2, 2022

The Mesquite mine is being operated by Gold Fields Operating Co.-Mesquite, a subsidiary of Gold Fields Mining Corp. The mine Is located in southeastern California about 56 km (35 miles) east of the farming community of Brawley, California. The mine has been in commercial production since early 1986 and Is a conventional open pit mine developed on 6.1 m (20 foot) benches. The gold is recovered by heap leach using permanent heaps built in multiple 6.1 m (20 foot) lifts. This year's gold production will be in the range of 5,600 kg (180,000 troy ounces). Current ore production is about 18,140 tonnes (20,000 tons) per day while the total tons mined Is averaging 65,000 tonnes (72,000 tons) per day. The climate at the mine is extremely arid with rainfall of 7.6 cm (3 inches) per year. From early May until late October daytime temperatures generally exceed 37.8 degrees Celsius (100 degrees Fahrenheit) and can be as high as 37.9 degrees Celsius (120 degrees Fahrenheit).

Jan 1, 1989

By Joseph A. Dos Santos

Developed in 1983, high angle conveyors - HAC®s are proven versatile systems for elevating materials continuously from a lower to an upper level. Such systems may also be used to lower materials continuously. HAC®s have found commer­cial application in widely varying materials (including coal, refuse, coarse ores [-250 mm], hot clinker, municipal sludge, wood chips, gypsum, slag, excavated earth, various grains and RDF), at widely varying throughput rates from 0.27 to 4250 t/h. The conveying profiles have varied widely and elevating heights range from 3.66 m to 175 m. HAC®s have found application in coal, copper and iron mining, and preparation, in municipal waste treatment, rapid transit projects, pulp and paper mills, cement plants and in gold mining. While noting a wide variety of applications in general, the present writing focuses on Newmont Gold's HAC®, near Carlin, Nevada, which elevates very coarse gold ore to the screening plant. Additionally, upcoming HAC® applications in gold mining are discussed and a proposed multi-flight vertical HAC® system from a deep underground mine is described.

Jan 1, 1996

By R. A. Basse, W. D. Diment, A. J. Petrina

INTRODUCTION In 1957, diamond drilling on a magnetic anomaly indicated an extensive zone of copper mineralization on what is now the Craigmont Mines property. By mid¬1958, drilling established a copper ore body. Milling commenced in September 1961 at 4536 t/d (5000 stpd) and by the end of October 1977 the mine had produced 339 662.04 t (374,363.9 st) of copper. At present, two-thirds of the mill feed is derived from underground operations and one-third from low-grade surface stockpiles. Craigmont Mines is situated 209 km (130 air miles) northeast of Vancouver (see Fig. 1), 16 km (10 miles) west of the town of Merritt, a logging, ranching, and mining community of about 7000 people. It is serviced by paved highways, Canadian Pacific Railway, British Columbia Hydro, and Inland Natural Gas Co. Water is pumped from the Nicola River, a distance of 6 km (4 miles) and a lift of 244 m (800 ft). In March 1967, the open pit mining operations at Craigmont Mines reached their economic limit and were suspended. Before this, it had been decided that a sub¬level caving method of underground mining would be used to supply ore to the concentrator after the cessation of open pit production. This chapter describes the fac¬tors influencing the choice of mining method, some of the problems encountered, mining practices, and results. GEOLOGY The ore bodies of upper Triassic age are located in a limy horizon striking east-west, closely paralleling the intrusive Guichon batholith, bounded on the south by rhyolites and on the north by graywackes, and dipping steeply to the south (Figs. 2a, b). The ore bodies are relatively narrow with a maxi¬mum width of 79 m (260 ft), a combined strike length of 853 m (2800 ft), and a vertical extent of 610 m (2000 ft). Chalcopyrite is virtually the only copper mineral, and 20% of the ore zone consists of acid solu¬ble magnetite and hematite. The area has been subjected to considerable faulting and brecciation, which is a major factor in the mining operation. Total geological reserves, at 0.7% Cu cutoff, for the deposit were 22 316 743 t (24,600,000 st) at 1.89% Cu. An additional 5 236 270 t (5,772,000 st) at 0.6% Cu were mined from the open pit. Ground Conditions The waste rocks-graywacke, andesites, and diorite -are relatively incompetent due to the high degree of fracturing and jointing, and all require varying degrees of support. The ore zones are somewhat less fractured; ground support is still required, however, although to a lesser extent than in the country rock. Ground conditions in the main ore body are better than in the smaller, nar¬rower ore bodies. Clayey fault gouge is present in most of the faults; gouge zones may be up to 6 or 9 m (20 or 30 ft) wide. The main ground problems are associated with local weakness rather than pressure. Shape of Ore Bodies (Figs. 2a, b and 3a, b) The main No. 1 ore body is approximately 244 m (800 ft) long and 46 m (150 ft) wide. It extends ver¬tically from the original top of the open pit at 4200 ele¬vation to just below the 3060 level. The No. 2 ore body is approximately 304 m (1000 ft) long, varies from stringer width at the extremities up to 79 m (260 ft) wide, and extends from 3060 level to 2400 level. Both these ore bodies have extensions re¬sulting in additional small irregular bodies. Ore bodies are mostly steep dipping, though part of the Wing ore body, an extension of No. 2 ore body, dips at 0.87 rad (50'). This ore body varies in size, but is approximately 122 m (400 ft) long, 21 m (70 ft) wide, and about 213 m (700 ft) high. No. 1 Limb ore body is a narrow extension of the No. I Main with a vertical extent of 137 m (450 ft), average width of 18 ft (60 ft), a strike length of 152 m (500 ft), and dips steeply at 1.4 rad (80°). No. 1 East is an eastern extension of the No. 1 Main with a vertical extent of 183 m (600 ft), a strike length of 91 m (300 ft), an average width of 30 m (100 ft), and dips at 1.2 to 1.4 rad (70 to 80°). No. 1 South is at the upper west end of the open pit with a vertical extent of 76 m (250 ft), a strike length

Jan 1, 1982