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[Participating Scholarship Colleges and Universities 136 Big Sky Coal Company 111 Canyon Fuel Company LLC 97 Coal Age Magazine 109 Commercial Testing & Eng. Co 99 Continental Conveyor & Equip. Co 103 Copper Mountain Resort 128 Dakota Coal Co 100 DBT America Inc 117 Empire Machinery 101 Energy West Mining Co 102]

Jan 1, 2001

By Sean Dessureault

Information technology (IT) initiatives in the past have tended to be sold on their ability to maximize resource output and return on capital by using assets more effectively and efficiently. Many IT projects fail, not delivering return on their investment. This is known in the IT industry as the ?productivity paradox,? where increasing investments in IT do not result in comparable increases in productivity. Experiences and studies of most business sectors reveal that this paradox is rooted in the false assumption that more IT spending automatically results in financial benefit, whereas the true source of benefits is in how effectively IT is used. Most business sectors have accepted that the reengineering of processes is required to gain the full benefits promised of IT. A mining operation has organizational processes similar to most business sectors but unique operational processes. The lessons of changing business practice in more common business processes such as inventory control and customer-supplier relationships can be adequately transposed into a mine. The unique operational processes in mines, in turn require unique redesign efforts. Several enabling information technology opportunities have recently become available, such as more integrated mine production management systems and tactical manger training simulators. This paper reviews the latest IT products and ou lines the processes and kinds of change that will be required in order to benefit from these new investment options.

May 1, 2002

By Roger L. Beals, George A. Tweedy

I. INTRODUCTION. THE results of the cyanide-operations, given in detail in the following paper, show the progress that is being made at the Minas del Tajo. The old pan-amalgamation process, in operation up to and including 1905, while paying a profit on high-grade ores, was no longer able to treat profitably the ores in sight at that time, hence an alteration in the process was imperative. The success of this alteration is eloquently expressed by the following data Pan- Amalgamation. Cyanidation. 1905. 1907. 1903. Tons treated, . . 51,387 53,983 69,300 Value per ton (U. S. currency),. $10.75 $8.39 $8.28 Extraction, per cent., 65.00 88.50 S7.45 Extraction, per ton. .- $7.00 $7.4.; $7.24 Total cost, per ton treated, 6.50 4.311 4.01 Profit, per ton treated, . . 0.50 3.07 3.23.;

Feb 1, 1910

By A. J. Barry

Recent developments in mechanizing longwall it mining in Europe have aroused considerable interest among American coal-mine operators because this method provides an opportunity to improve safety, production, efficiency, and recovery. The success of longwall mining in Europe has depended to a large degree on development of effective roof-control systems. One of the major contributing factors in designing these systems is development off improved supports, which combine high load-bearing capacities with controlled yield characteristics, to use in the face areas. Other factors that relate to successful longwall mining involve knowledge of the structural properties of the strata above end below the coal bed being mined, movements of these strata induced by mining, and effects of the strata movements on the strata itself, as well as on the support system used. Extensive studies of the above factors relating to long-wall mining have been made in coal-mining countries abroad, and corresponding studies (1, 2, 3, 4, 5)3/ of room-and-pillar operations have been made in this country. Although descriptions and results of experiments of mechanized longwall operations in the United States have been published (6, 7, 8), this is the first instrumented study of roof behavior on a longwall operation. In view of the possible widespread application of this method of mining to American coal mines, where underground mining conditions are different from those found in European mines, an investigation was conducted at a mechanized longwall operation inn the Lancashire No. 15 mine, Barnes & Tucker Co., Cambric County, Bakerton, Pa., to study the factors that affect roof control. This report describes this investigation and discusses the results. ACKNOWLEDGMENTS This investigation was made under a cooperative agreement between the Bureau of Mines and the Barnes & Tucker Co., Barnesboro, Pa., end was conducted in the Lancashire No. 15 mine located at Bakerton, Cambria County, Pa. The authors acknowledge the cooperation given by Richard Z. Todhunter, Sr., president, Richard T. Todhunter, Jr., general manager, and other mine officials and employees of the Barnes & Tucker Co. Mr. Edward Thomas, chief, Roof Control Section, Bureau of Mines, Washington, D. C., and Dr. Leonard Obert, chief, Applied Physics Branch. College Park, Md., collaborated in the planning and promulgation of the program investigation.

Jan 1, 1954

By G. G. Krishna Murthy

Several experiments were carried out to simulate the bath smelting reactions where iron oxide particles were dropped on to an iron-carbon bath containing molten CaO-SiO2-Al2O3 slag layer on its surface. The composition of the slags used in the experiments was %CaO:46.1; %SiO2:41.7; and %A1203:12.2. Experiments were con¬ducted with carbon saturated iron melts contained in either a pure graphite crucible whose walls were covered with a pure alumina tube, or a pure flat bottom alumina crucible, at 1723-1823 K. Degree of reduction of a pure iron oxide particle was linear with time in the initial stages (up to about 20% oxygen of the original pellet was reduced). The rate of reduction of smaller oxide pellets was smaller than that of a bigger pellet - the time for complete reduction of a 5.6 g wustite pellet was smaller than that for a 1.2 g wustite pellet. It was also established that the interaction rates obtained by adding a molten pure oxide particle of given weight to the system were lower than that determined by using a solid oxide particle of same weight. The overall rates of reduction decreased with decreasing the bath temperature. The experimentally determined rates of reduction based on the time for 70% reduction were influenced substantially by addition of slag up to about 50 g on to a bath of 78.8 mm in diameter, and remained constant when slag present on the bath surface was between 50 g and 200 g.

Jan 1, 1993

By Robert D. Harris

In 1958, the coal industry, through Bituminous Coal Research, Inc., and the electric utility industry, through the Edison Electric Institute and the Association of Edison Illuminating Companies, initiated a joint program to study coal pyrite and its removal as a means of lowering the sulfur content of steam coals. It was assumed that significant amounts of pyrite often remained in washed steam coal not because available systems for coal cleaning were inadequate, but because the pyrite was small and widely disseminated, During the first two years, samples of coal were examined petrographically to determine the size, association, and mode of occurrence of the pyrite in the raw coal and in the products of laboratory separation, Samples of coal were also pulverized in different equipment to determine how effectively pyrite was released by pulverization and if the mode of pulverization affected the liberation of pyrite. From these studies, the following tentative conclusions were made (1 ): 1. Much of the coal pyrite is within a 250-70 micron size range, and therefore is too small to be removed by conventional coal cleaning, but relatively large when compared to the size of coal as pulverized for combustion in large boilers. 2. The degree of pyrite liberation is mainly determined by the size of the product and not by the type of crusher. 3. For improved pyrite removal, the coal should be cleaned during or after pulverization at the power station.

Jan 1, 1965

By D. G. F. Hedley, F. Grant

"In 1958, the Ontario Department of Mines appointed a committee to study the accident situation and mining methods used in the uranium mines in the E1liot Lake district. This committee reviewed the existing knowledge on mine design, pillar support and roof spans and came to the conclusion that scientific knowledge had not yet reached the stage of producing rational design procedures which would reduce the dependence on trial-and-error methods. Since that time, the Mining Research Centre has made numerous stress measurements in two mines to define the pre-mining and pillar stresses. A comprehensive series of tests has been done in the laboratory to measure the compressive strength of the rock; a geological survey of the region has been made to evaluate tectonic stresses; and a survey was done on mining conditions with regard to stable and unstable pillar configurations. This information has been used to evaluate pillar design as effected by pillar width and height, depth below surface, extraction and pre-mining stress fields. The resultant design of pillar widths and corresponding extraction as the depth increased is in line with past and present experience in these mines. The information available on roof stability is still insufficient to design stable stope spans with the same degree of confidence; experience and trial-and-error methods are still required."

Jan 1, 1972

By S. L. Bessinger

Introduction In the design and specification of a ventilation system for an underground mine, it is necessary to make reasonably accurate estimates of the pressure losses in the various airways of the mine. These estimates can be made with little difficulty for open airways with simple geometric cross-sections, such as those cut by continuous miners or tunnel-boring machines. The situation is much different on a longwall face, where the airway's complex geometric cross section and the presence in the airway of obstructing equipment having a variety of shapes make it difficult, if not impossible, to estimate pressure loss using traditional methods of calculation. Head losses in mine entries are calculated using Atkinson's Equation. [22H= KPLQ (English) H= KP 3Q (SO (1) 5.2AA] where H = pressure loss, in. of H2O (Pa); K = friction factor, lbf•min2/ft4 (kg/m3); P = perimeter, ft (m); L = airway length, ft (m); Q = airflow quantity, ft3/min (m3/sec); and A = flow cross-sectional area, ft2 (m2) In this equation, the friction factor, K, is an empirical constant that describes the aerodynamic roughness of the airway. Typically, the K-factor for a given airway is determined by measuring the factors H, P, L, Q and A in Equation (1) and calculating K. Tables of friction factors calculated in this way are found in textbooks and handbooks that deal with mine ventilation analysis. Unfortunately, very few K-factors have been measured on longwall faces, and the accuracies of those that have been measured are entirely site specific, because of the wide variety of equipment found on longwalls. The development of a technique for prediction without mine-site measurements of the friction factor for any longwall face, regardless of its configuration, will thus be very useful in the design of ventilation systems for mines in which longwall mining is practiced. Calculation of pressure losses using Atkinson's Equation (1) and empirically determined K-factors provides accurate and useful approximations in cases where the airways have relatively simple cross sections. However, a careful analysis using the principles of fluid mechanics shows that such calculations are based on two assumptions that are not strictly correct when there are obstructions in the airway. The first assumption is that the air velocity distribution in the cross section, particularly around the perimeter, is uniform. This assumption results from the fact that the tabulated K-factor values found in the literature are based on field measurements with uniform conditions. Such uniformity does not exist in longwall airflows. The second common assumption is that the K-factor, and corresponding head loss, is independent of the Reynolds Number (NR) for a given flow. In fact, this assumption is not strictly correct, and is particularly erroneous in the case of irregular protuberances into the airflow, such as those found on a longwall face. The errors arising from the assumptions may be avoided by using K-factors calculated by a newly devised method, described below, which takes into account the fundamental principles of aerodynamic drag analysis. This new technique has two advantages: first, it is flexible enough to model any longwall, regardless of equipment configuration; second, it employs terminology and equations familiar to those who perform mine ventilation analysis, using K-factors, for which ventilation engineers have an intuitive understanding, rather than drag coefficients. To provide guidance for development of a longwall drag model, data were taken on two modern longwalls operating in substantially different conditions. Pressure measurements at Mine B were made with 200-foot (61-m) sections of 1/8-in. (3-mm) diameter plastic tubing, attached to a Dwyer Magnehelic gauge. Pressure drops were measured in 200-foot increments down the face, and summed to give the drop for the entire face length. This technique was found to produce small, repeated errors because of the number of segments required to span the longwall. At Mine A this problem was avoided by using a single, continuous, plastic tube for the entire face length. The psychrometric properties of the air were measured for both Mines A and B. A calibrated, standard vane-anemometer was used to measure the airflow on both faces. Finally, numerous dimensions were measured on both faces, and face profile drawings were obtained to allow detailed evaluation of the face equipment geometry. From this information, accurate evaluations of the average wetted perimeter and average area of the longwall face airways were made. Since the airflow is not confined to inside the powered supports at all points along the face, a quadratically weighted average of the airflows measured at various stations along the face was calculated: [n2Qavg =Qi Ii / It(2)i=1] where [Q, avg = average airflow for analytical purposes, ft3/min (m3/sec); Q= airflow at station i, ft3/min (m3/sec); 1= length of segment represented by Q, ft (m); h= length of longwall face, ft (m): and n= number of quantity measurement stations.] The quadratic weighting scheme was chosen because the

Jan 1, 1992

By R J. Soriano, S Ciutina

The Masbate gold deposits are situated on the island of Masbate, 350 km south of Manila. A new plant was commissioned in April 2009. The grinding circuit consists of one semi-autogenous (SAG) mill Ø9.75 m × 4.27 m, two ball mills Ø5.6 m × 7.1 m and a pebble crushing circuit, which is usually called a SAG-ball mill-crusher configuration. The SAG mill was originally equipped with radial pulp lifters. The mill runs at fixed speed of 10.9 rev/min, which is approximately 80 per cent of critical speed (per cent CS). Due to the high speed of the mill the original radial design pulp lifters presented a high volume of recirculating load inside the chamber and consequently heavy wear on the outer circle (bottom pan).In order to address the high wear on the pulp lifters a number of avenues were considered in early 2011. The chosen solution was to change the geometry of pulp lifters from radial to curved design. The pre-existing drilling pattern on the discharge end of the mill was a major constraint. A number of designs were considered. The selection of the most suitable option was to be made based on 3D discrete element modelling (DEM) and smoothed particle hydrodynamic (SPH) simulations.The original radial design and two potential solutions for curved pulp lifters were simulated with 3D DEM and SPH software. The main three objectives were predicting the charge dynamics, the expected wear patterns and discharge efficiency. One of the solutions was chosen and implemented during 2012.Observations carried on after installation of the curved pulp lifters have confirmed, as expected, a reduction in wear intensity on the outer pulp lifters. In addition to this, at early stages of operation with curved pulp lifters, the mill operator noticed a reduction in energy consumed by the mill with no detrimental effect on mill performance.This paper is intended to further study the effect of curved pulp lifters on mill energy consumption.CITATION:Ciutina, S and Soriano, R J, 2014. Curved pulp lifters – can they save energy?, in Proceedings 12th AusIMM Mill Operators’ Conference 2014 , pp 109–116 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Sep 1, 2014

By Edgar Peek

"The two dominant metallurgical problems in treating most base metal sulphide concentrates in both the chloride and sulphate system are iron removal and sulphur elimination, respectively [1-3]. In this paper the opportunities, limitations and fundamentals of the chloride metallurgy system are described against this background. Only the treatment of oxide and sulphide ores is considered. Both hydrometallurgical and pyrometallurgical unit process operations are highlighted, since most successful base metal processes are based on a combination of the two. Hence, an overview of the technical and also some economical aspects of chloride metallurgy are presented.This overview summarizes the successful applications of chloride metallurgy, but it will not give exhaustive process and plant descriptions. It predominantly focuses on the essential technical features of the metallurgical unit process operations, while providing numerous references separately."

Jan 1, 2011

By Robert S. Shoemaker

The development of a metallurgical project is a business venture rather than a technical exercise, and its real objective is to obtain the maximum return on investment consistent with the limitations imposed by company policy and the requirements of society. Demand for profits is the driving force for technical advances and naturally leads to improvements in processes, equipment and plant designs. In the mineral dressing field there have been virtually no new processes developed since the introduction of flotation, and the steady erosion of profits by inflation has been offset only by increasing equipment size. Unfortunately, in recent years, plant designers have not always utilized the largest equipment or the opportunities for functional and economic design presented by the development of larger equipment. Lower cost plants can be developed providing that unrealistic restraints arising from the prejudice of past practices are avoided. As an illustration of this premise, three different approaches were taken toward the design of a clearly delineated but hypothetical copper concentrator of 20, 000 tons per day capacity. While the same flowsheet was used in each case, designs were produced for a conventional plant, an alternate using the largest available equipment, and a third scheme utilizing large equipment and the most functional design which could be developed, including features derived from many plants in many different industries. Capital costs were developed for the three separate designs based on a location in the southwestern United States. Ancillary facilities were identical in each case and included a mine truck shop, a maintenance shop, a general office, and a nominal starter dam for tailings disposal within 5000 feet of the plant. Water supply within 1000 feet of the plant, and power furnished to the property boundary has been assumed in each case. The flowsheet of the three plants is shown in Figure 1 and represents the treatment of a typical sulfide copper ore by three- stage crushing followed by single-stage ball milling. Concentration is by flotation roughing and scavenging followed by regrinding of the concentrates and two stages of cleaning.

Jan 1, 1971

By Wei Luo

Nickeliferous laterites were characterized by particle size analysis, X-ray fluorescence (XRF), X-ray diffraction (XRD), scanning electron microscope (SEM) and energy dispersive X-ray analysis (EDX), Fourier transform infrared (FTIR) spectroscopy and thermal analysis. Results showed that the laterites consist mainly of nickel-substituted lizardite with cobble-like morphology and traces amounts of magnetite and phlogopite. As fluoride ions can react with silicon in the lizardite crystal structure and form metal-fluoride complexes, as well as increase the equilibrium constant of the lizardite dissolution in acid solution dramatically, the addition of minor fluoride salts (0.2wt.% NaF) gave a markedly enhanced metal extraction rate (Ni 81.7% and Co 52.3%) under the experi-mental conditions of particle size d50= 25 µm, 10% (v/v) H2SO4, reaction time 2 h, reaction temperature 90°C, liquid-to-solid ratio 3 mL g-1 and stirring at 500 rpm compared with H2SO4 alone (Ni 69.6% and Co 30.8%).

Jan 1, 2010

By Chu E. Ho, Alfredas Daugiala, Nicola Ebanks

"The construction of an ancillary structure for the new Second Avenue Subway Line in Manhattan, New York City involved a deep excavation in highly variable rock conditions adjacent to a fragile masonry building. The maximum difference between the lowest and highest rock elevations was about 30 feet, which posed a problem for the installation of secant piles that formed the supporting retaining wall for the excavation. At certain areas, the rock levels were higher than the top of the invert slab of the ancillary structure. To avoid disturbance to the adjacent fragile building, the secant piles were fully cased for drilling through the overburden soils and the rock sockets were formed using non-vibratory technique. This paper discusses the design and installation details of the secant pile wall and the contingency measures that were considered to mitigate the highly variable rock conditions. Monitoring of the adjacent fragile building showed that impact of pile installation were controlled within tolerable limits and safety of the building was not compromised.INTRODUCTIONThe 96th Street Station forms part of the newly constructed Second Avenue Subway Line on the East side of Manhattan Island, New York City (Figure 1). During construction of this station, several deep excavations were carried out adjacent to existing fragile buildings (Arksman et al. 2012, Grigson et al. 2016). These fragile buildings were typically four to five stories high, with a single rubble wall basement supported on timber piles or seated directly on the fill. The timber piles were often found to be in a poor condition and connections to the footings were compromised due to rotting of the timber within the zone of water table fluctuation. Several of these fragile buildings had to be strengthened and underpinned to improve the stability of the structures prior to commencement of excavation for the new subway station box (Trabold et al. 2012, Ho and Pena-Iguaran, 2013). The situation was most challenging at Ancillary Buildings 1 and 2 located at 93rd and 97th Street respectively, where secant pile walls were located merely 2 feet away from the adjacent property line, due to architectural space constraints. While mitigation of excavation induced ground movements can be controlled by the adoption of a stiff excavation support system, a more concerning factor was the uncertainty associated with the impact of the secant pile wall installation on the stability of the existing fragile buildings. This was particularly difficult at Ancillary 1 where the secant piles encountered highly variable rock levels that intruded into the excavation space. The following sections discuss the design and installation of the secant pile wall and contingency measures that were considered at Ancillary 1."

Jan 1, 2016

By H. H. Schrenk

One phase of Bureau of Mines work in health and safety, is promotion of the use of safe, satisfactory, and reliable respiratory protective devices. To insure that such devices will be available, the Bureau has established schedules of minimum requirements for self-contained breathing apparatus,3/ gas masks, supplied-air respirators (hose masks, air-line respirators, and abrasive blasting helmets, hoods, or masks),2 dispersoid (dust, fume, and mist) respirators, and nonemergency gas respirators (chemical cartridge respirators). Approval tests of self-contained breathing apparatus are conducted by the Safety Branch, approval tests of other types of respira- tory protective equipment are conducted by the Health Branch, These branches are in the Health and Safety Division.

Mar 1, 1948

By W. Rogers Wade, Alfred Wandtke

The Pilares mine of the Moctezuma Copper Co. is situated at Los Pilares de Nacozari, Sonora, Mexico, about 75 mi. (120.7 km.) south of the international boundary and about 7 mi. (11.26 km.) east of the town of Nacozari de Garcia. It is situated at an elevation of 5000 ft. (1524 m.) in one of the northward trending ridges that form part of the mountain mass that culminates in the Sierra Madre de Mexico. To the east, south, and west, the topography is rugged and diversified; to the north less diversity exists, ranges of mountains rise sharply out of almost level ground and give the basin range structure characteristic of the Southwest. The climate is dry, the average rainfall of 15 in. (38 cm.) being confined almost entirely to the months of June, July, and August. The winter consists of a few days of disagreeable weather when a little snow may fall; the remainder of the year is thoroughly delightful. Vegetation is typical of a semi-arid region. In the canons near the mine, where water flows intermittently, willows thrive; on the northern slopes of some of the mountains and in some of the broader canons pines may even become abundant. The dominant rocks are flows and pyroclastics which are cut by dykes, stocks, and batholiths. The mountain just west of Nacozari is, in part, composed of limestone of an unknown age which occurs as a syncline and is undoubtedly a part of a prevolcanic land surface. Overlying the limestone and faulted against it, are the volcanics; cutting the limestone is a coarse biotite granite. Whether the granite is older or younger than the volcanics has not been definitely established. No attempt has been made to differentiate the rock types of the volcanic~, of which there are exposed 3000 ft. (914 m.) in thickness of basic and acid flows (andesites, latites, and rhyolites) and andesite and latite tuffs and breccias. The latite breccias, the youngest of these volcanic~, cover the highest mountains in the immediate vicinity. The

Jan 1, 1920

By D. D. Munro

Finsch Mine is a Kimberlite diamond mine located at Lime Acres in the Northern Cape of South Africa. The mine was founded in 1961 and started surface-mining in 1964. Underground production commenced in 1990 using a Modified Blast Hole Open Stoping method for the mining of Blocks 1,2 and 3. Block 4 is currently being mined as a Block Cave. The process of identifying and optimizing a mining method to mine the Block 5 ore body started in 1991 and in 2006 Inclined Caving was identified as being a technically feasible method. This report aims to document the process employed in developing this method by the Block 5 Pre-Feasibility Team as well as discussing the technical challenges encountered during this process.

Jan 1, 2010

By Jeffrey Y. Foley

In 1980, the Bureau of Mines investigated the northern Darby Mountains on the Seward Peninsula, AK, for radioactive mineral deposits. Uranium is concentrated in silicified shear zones in biotite quartz monzonite of the Darby pluton. The shear zones are best exposed in steep, narrow, avalanche gullies; they are characterized by radiometric anomalies and by uranium and thorium geochemical anomalies. Radiometric and geochemical anomalies are also found over altered biotite quartz monzonite where there are no visible shear zones. The radiometric and uranium anomalies are often spatially, but apparently not genetically, associated with alkaline mafic and felsic dikes. Similar geochemical and radiometric anomalies over discolored altered zones and geochemical anomalies in sediments downstream from the altered zones indicate that additional areas may also be favorable targets for uranium deposits. Analyses of panned, heavy mineral concentrates show no significant concentrations of uranium or thorium in resistant heavy minerals. This indicates that radioactive elements in the Darby pluton are concentrated in soluble, nonresistant minerals that may constitute a recoverable uranium resource.

Jan 1, 1986

By Gordon M. Pugh, David G. Rasmussen

INTRODUCTION Sublevel long-hole stoping is a large tonnage stoping method that may be used when there is a fairly com¬petent straight vein and the walls will stand with little support. It is a method that lends itself to a high degree of mechanization. Two variations of the method are illustrated in Figs. 1 and 2. Fig. 1 illustrates development typical for a 3-m (10-ft) wide vein. Initial development is by a lower haulageway, which may be either track or trackless. A two-compartment raise is driven in the footwall of the vein. At about 12 m (40 ft) above the main head¬ing, a footwall scram is driven and pockets through the vein are cut on 7.6-m (25-ft) centers. Each of the sub¬level blocks selected in this example is 45.7 m (150 ft) long by 12 m (40 ft) high. Stoping starts at the slot raises at stope endlines and retreats to the timbered raise. This retreat sequence starts on the lowest sublevel and the opening cut is blasted into the slot raise. This retreating system is used in order to provide a good safe back over the miners working in any given sublevel. A crawler-mounted down-the-hole or conventional pneumatic drill puts the holes down from one sublevel to the next on any desired drill pattern. Gravity flow brings the broken ore down to the scram level. The illustration shows a large slusher mounted at the timbered raise and slushing from the pockets back to the raise chute. Development of the ore block in Fig. 2 is typical for a 5.5-m (18-ft) wide vein. Initial development is by a trackless heading driven in the footwall of the vein. Pockets are then put into and through the vein on 7.6-m (25-ft) centers. For ease in using trackless rubber tired mucking equipment, these pockets are crosscut to the vein at approximately 1.04 rad (60°) rather than per¬pendicular. A two-compartment timbered raise is driven down from one level to the next. Endline slot raises are again put through from level to level. Each of the sub¬level blocks selected in this example is 40-m (130 ft) long by 33.5 m (110 ft) high. Stoping starts at the slot raises and retreats back to the timbered raise. Long holes are drilled from one sub¬level down to the next by starting between the lowest pair of sublevels and then working up the ore block. Mucking out of the pocket drawpoints is accomplished by rubber-tired load-haul-dump (LHD) units which then dump into a centrally located orepass.

Jan 1, 1982

By R. C. Nascimento

It was investigated the effect of the addition of bentonite on the physical propertiesof pellets produced from fines of manganese ores. Cold-bonded pellets were produced. Pellets bearing up to 1 wt% of bentonite were submitted to strength compression resistance tests, fatigue of pellets to impact and maximum height of fall tests. A correlation of resistance of the pellets and time for curing were stablished. The produced pellets match the conditions for raw materials used in electric smelting hrnaces to produce manganese based ferro-alloys.

Jan 1, 1995

By M. Bryson, J. Taguta

The Prieska deposit, mined between 1971 and 1991, is one of the world’s top thirty volcanogenic massive sulphide (VMS) base metal deposits. The Prieska Copper Mine in the Northern Cape Province was closed by Anglovaal when underground mining had extended to depths which did not allow for safe and economic extraction of the deposit at the time. The mine has a recorded historical production of over 1 Mt of zinc and 430 k tonnes of copper from 46.8 Mt of milled sulphide ore, achieving life-ofmine metal recoveries of 85% for copper and 84% for zinc into high quality concentrates (refer ORN ASX release 15 Nov 2019). Orion Minerals Ltd is currently investigating the feasibility of developing a new mine in the footprint of the historical Prieska Copper Mine, using the extensive remaining infrastructure, with the intent of extracting the unmined zinc-copper mineralisation from the Prieska deposit. This paper reports results on the flowsheet development philosophy on the hypogene ore, and challenges encountered in the process. This study initially failed to replicate historic copper and zinc recoveries employing the historic production flowsheet and reagent conditions unmodified on bench scale, as the copper and zinc were co-floating. Flowsheet modifications were therefore made to improve differential flotation and recovery, resulting in achieving comparable saleable products and recoveries from processing the hypogene ore on bench scale. Flotation variability test work on the hypogene ore focused on the use of the original flotation procedure and reagent suite using a sodium cyanide-zinc sulphate depressant system on the copper circuit, and it generated a saleable copper product of 20-26% Cu grade at 80-86% recovery. The zinc circuit generated a saleable zinc product grading at 48-55% Zn at 80-94% recoveries. Some zinc losses were due to the deportment of some zinc in sauconite. It is recommended that further test work may further improve the copper grades and recoveries. Keywords: Prieska, flotation, hypogene ore, flowsheet development, sulphide flotation, reagents

Jan 1, 2020