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By M. Lonoff

G. Campbell The paper by M. Lonoff looks at the importance of byproduct prices on copper production. The paper develops several interesting points on this topic, but there are some points in the theoretical discussion that could use further development. A couple of these theoretical topics will be considered here using the paper's framework of assumptions without concern about how realistic the assumptions are for the copper industry. Lonoff's theoretical model is based on a static, homogenous world where all copper/byproduct deposits are identical, and these deposits are the main source of these metals. Due to the importance of the byproducts to the production decision, the term coproducts will be substituted for the term byproducts to reflect more accurately this condition. The conclusion is drawn that a rise in coproduct prices necessitates a decrease in copper prices because copper production is increased as a consequence of increased coproduct production. The observation that real world behavior does not always follow this pattern is dismissed as "speculation." However, it can be easily shown that within this simple framework market forces might create the observed behavior. The key question to be answered is what causes coproduct prices to rise. Three possible cases will be used as illustrations. The first case is the one implicity assumed by Lonoff. Here, the demand for coproducts increases at all prices (an outward parallel shift of the demand curve) with the demand for copper remaining constant. As reported in the paper, the results are higher coproduct prices, an increase in coproduct production, and lower copper prices due to the resulting increase in copper production. The second case is when the supply of coproducts are constrained-leading to higher market prices. Copper production, as required by the simple model, is also reduced which Reply by M. Lonoff Campbell's comments give me an opportunity to clarify a few points in "Copper and Its Byproducts." That paper examined the affect of byproducts on the copper market and considered the relative riskiness of multi-metal deposits from an investment standpoint. That brief note made no attempt to be exhaustive. Nevertheless, Campbell's comments miss the point of the paper and necessitate a response. My "theoretical model" was static and I freely conceded after introducing the assumption of identical deposits that "none of the assumptions hold." These simplifying assumptions allowed me to bound the possible effects of higher byproduct prices on the copper price by assuming identical deposits and computing average price effects. I took great pains to subsequently explain that since deposits are not identical the identity of the marginal deposit might change due to byproduct price changes. In this case, average price affects will overstate the affect of byproducts on copper prices. According to Campbell, I dismiss as speculation any "real world behavior that does not always follow this pattern" (of metals prices moving in opposite directions). In the byproducts paper I noted that gold, copper, and silver price movements often parallel each other on an intra-day or day-to-day basis. Such movements are largely due to market anticipation of or reaction to changes in interest rates that affect carrying costs of metals stocks. Other movements may be due to anticipated changes in available supply or physical requirements. I loosely grouped these effects as speculative. To distinguish this statement from the main thrust of the paper, I contined "In the long run if gold and silver prices remain high the equilibrium copper price will be lower." The paper set out long run equilibrium conditions rather than attempting to explain price movements on a daily basis or on a year-to-year basis.

Jan 11, 1983

By J. A. Marshall

"INTRODUCTION The purpose of this paper is to investigate the potential for use of UAVs in underground mines and present a prototype design for a novel autorotating UAV platform for underground 3D data collection. The mining industry has recently shown increased interest in the use of unmanned aerial vehicles (UAVs) to assist in everyday operations [20, 11, 4]. Above ground, small UAVs are in some cases a more efficient, inexpensive, and safer alternative to manned aircrafts currently used for photography, inspection and security [6, 10]. For example, attaching camera, infrared and LiDAR payloads, UAVs can provide a low-cost method of obtaining highly accurate 3D photogrammetric data and aerial photography. UAVs are now commonly used in open pit mining operations for applications that include stockpile surveying, 3D pit modelling, facilities management, accident reporting, progress monitoring, and environmental assessment [11]. There are numerous companies offering UAV services for mining applications; e.g., UAV Geomatics, Leica Geosystems, Sky Futures, SenseFly and FlyTerra, to name just a few. Common services include stockpile and open pit volume computations, environmental assessment, aerial mapping and photogrammetry. Although the aforementioned services are thosemost commonly provided by UAV surveying companies, they may also add value to other daily mining operations. For instance, SenseFly offers pre and post-blast monitoring in order to identify the presence of misfires and wall damage. This data can also be used to reconcile the blast results with expected results. GEM Systems offers a UAV equipped with a magnetometer for mineral exploration surveys. Companies such as Barrick have even used UAVs for solids modelling at tailings dams and stability monitoring [11]. UAVs in mining have so far been mostly limited to surface applications. Harsh underground environments pose many obstacles for flying UAVs. The confined space, dampness, reduced visibility, air movement, and lack of control signal propagation hinders most operators from being able to fly a drone underground. It may be that truly practical uses for UAVs underground will require either autonomous or semi-autonomous flight capabilities. Although there are many difficulties with flying underground, the potential benefits from a working system could greatly improve mining operations. The potential benefits of deploying UAV platforms underground include access to unreachable and dangerous locations and removing personnel from unsafe situations. These benefits have the potential to greatly improve mine monitoring and mine safety. Research has shown that current UAV technologies exist that allow for autonomous indoor flight. Extensive research has been done to develop UAV systems that are capable of performing on-board simultaneous localization and mapping (SLAM), which can allow them to navigate and map a foreign environment autonomously [5, 7, 1, 13]. Grzonka [7] successfully used an open hardware quadrotor to autonomously navigate and map an office building. The research outlines the localization, mapping, path planning, height estimation and control of the autonomous quadrotor. Other research has been done that exploits autonomous UAVs for search and rescue. Kassecker [9] proposed a software and hardware framework for a quadrotor capable of indoor and outdoor urban search and rescue and Rudol [19] developed a system for human body detection and geolocalization using an autonomous UAV.The use of autonomous UAVs in search and rescue has the potential to improve situational awareness and surveillance for a rescue team. Similar research has been done that uses autonomous UAVs for indoor exploration. Pravitra [15] outlined a strategy for autonomous exploration of miniature air vehicles (MAVs) within indoor environments and Rudol [19] developed a method for using an unmanned ground vehicle (UGV) and a UAV for cooperative indoor exploration. These studies show that aut"

Jan 1, 2017

By Richard R. Klimpel

INTRODUCTION Froth flotation is the most widely used and economic means of concentrating metal sulfide ores such as those containing copper, lead, zinc, nickel, molybdenum, and pyrite. Also recoverable are other metal species that are often associated with sulfide ores such as cobalt, platinum, gold, silver, etc. Froth flotation is a physico-chemical process for separating finely ground minerals from their associated gangue material. The process involves chemical treatment of a finely divided (ground) ore in a water pulp to create conditions favorable for the attachment of certain of the mineral particles to air bubbles. The air bubbles then carry the selected minerals, called valuable, to the surface of the pulp to form a stabilized froth which is removed and recovered. The unattached gangue material remains submerged in the pulp and is either discarded or reprocessed. To obtain the adherence of the desired mineral particles to the air bubbles, at least two specific steps must occur: a hydrophobic (water hating) surface film must be formed on the particles to be floated along with a hydrophillic (water wetting) film on all other particles; and a controlled bubble surface tension interface must be maintained, allowing for high particlelbubble collision frequency and efficient attachment or sticking of the particle to the bubble once collision has taken place. In most flotation applications, the above two steps are controlled by chemical flotation reagents. The collector is a chemical reagent which produces the hydrophobic film on the valuable mineral particle and is the primary driving force that initiates the flotation process. The frother is a chemical reagent which influences the collision frequency and attachment efficiency of hydrophobic particles and air bubbles. Thousands of research papers and books have been published on the chemical theory behind sulfide mineral collectors, e.g., Fuerstenau (1962), Fuerstenau (1976), Fuerstenau, et. al. (1985), King (1982), Leja (1982) and Moudgil and Somasundaran (1987), This article will only address and summarize some of the more practical aspects of collector usage. The industrial scale practice of froth flotation in sulfide mineral concentration has changed little since the 1950's. For example, of the approximately 80,000 metric tons of sulfide mineral collectors used commercially (1980) in the free world, almost 98% were known and manufactured in some form 25 years ago. A very interesting and informative history of collector development has been given by Crozier (1984). In addition, the industrial scale practice of froth flotation applied to sulfide ores has proceeded since the 1920's with often little direct (predictive) scientific explanation due to the extreme complexity of the flotation process. Empirical testing has been a mainstay of industrial flotation reagent development and use. Even today there is often strong disagreement between researchers as to the mechanisms of chemical flotation practices that have been performed successfully at an industrial scale for many years. As a result of the above environment which makes reagent cost/performance analysis difficult for new reagents, there is a strong tendency for the flotation operation to use, and reagent companies to supply, as cheap as possible raw materials and manufactured products that are quite general in application. In the last 20 years or so, there has been increasing technical effort to tailor make reagents for specific applications, but to date such work has had little commercial impact. It is clearly the hope of this author and others that this situation will change in future years as technology improves and pressure for improved flotation performance intensifies. This article is a condensation of collector usage trends quantified as part of a comprehensive industrially oriented applications program on froth flotation organized by Klimpel and coworkers (1979-1987) and as reported in various countries that participated in the program from 1978-1983. No attempt will be made to provide specific.

Jan 1, 1986

By P. D. Chamberlain

Agglomeration for the heap leaching of gold and silver ores is the process of attaching fine ore particles-less than 150 µm (100 mesh)-to coarser particles. The fines are thus immobilized and uniformly distributed throughout the heap. This is needed to make heaps more porous and uniformly permeable to the flow of a leaching solution. This article describes liquids and binders used to form agglomerates. Agglomeration techniques and equipment are also described. These include drum and disk agglomerators, stockpile and belt agglomeration, and the use of vibrating deck and reverse-running conveyors. The need for agglomeration The main reason for agglomerating heap leachable ores is to prevent percolation problems in a heap. This is different than pelletizing iron ore. In the latter case, the purpose is to bond very finely ground material into coarser, strong particles. These particles can then withstand the rigors of transportation and the loading forces in a blast furnace. Most percolation problems in heaps are caused by the segregation of coarse and fine particles during heap construction. This segregation creates areas with significantly lower permeability because fines plug the channels between coarse particles. Consequently, the leach solutions are subject to channeling, following the path of least resistance. They percolate downward through the coarse ore areas and bypass or barely wet areas that contain large amounts of fines. This results in lower extraction, longer leach time, and higher reagent consumption. Permeability problems in heap leaching are compounded because of the vertical migration of fine particles after the heap has been built. These fines eventually accumulate in pockets or layers and impede uniform solution flow. The fines move because they are not attached to a coarser particle. Moreover, the sluicing action of leach solution flowing over the surface of particles may break the bond between fine and coarse particles. Some fines are also created by the decomposition of ores after being wetted. Percolation problems are minimized if the fines can be attached to the coarser particles. Fines are thus uniformly distributed and rendered immobile within the heap. Heinen (1979) conducted comparative tests of nonagglomerated and agglomerated gold and silver ores. He showed that percolation rates through agglomerated material usually are improved by 10 to 100 times. This reduces leach time for economic recovery to less than one-third of that for nonagglomerated material, and total recovery may be increased. Cyanide consumption is also reduced in most cases because of the shortened leach time. Even run of mine (ROM) ore should be agglomerated. This can be achieved with moderate success by passing each truckload of ore under a water spray to bring total moisture of the ore up to about 8% to 10%. The ROM ore will mix when the ore is dumped, when it cascades down the slope of a heap, or when it is pushed up with a dozer. Overall, agglomeration allows low-cost heap leaching technology be applied to gold and silver ores and tailings that could not be economically processed by conventional methods, according to Phariss (1982). How agglomeration works Agglomeration means that the clays and fine particles contained in the ore adhere to the coarser particles. This creates a coating of fines around the coarse particles. When the ore has few fines and the percolation problems are not severe, the addition of a liquid may be all that is necessary to make the few fines adhere to coarser particles. However, when the quantity of fines is substantial, e.g., +10% -75 µm (-200 mesh), a binding agent may be needed. Use of solution only for agglomeration There are many forces that cause agglomeration and one of them is the surface tension of water. Water-saturated particles collide with each other or with

Jan 12, 1986

By Robert D. Hansen, Richard R. Klimpel

INTRODUCTION A very important factor in froth flotation is the use of surface-active chemicals to form a froth in which minerals or coals are retained, thus allowing for valuable component enrichment. In coal flotation reagents (frothers) specifically designed to provide an efficient froth are employed. It is this group of chemicals that will be the primary emphasis of this paper. Despite the importance of frothers, surprisingly little systematic scientific work has been performed with the goal of quantifying frother mechanisms and the influence of particles on froth stability. Based on direct experimental observations in both batch laboratory and continuous large-scale flotation cells, the use of a frother significantly increases first, the possibility of a particle- bubble contact, and second, the efficiency of sticking after such a contact. Thus, a major role of a frother is to significantly increase the rate of flotation. This rate modification by a surface-active agent occurs through several observable mechanisms including the formation of a froth of relatively consistent character (bubble size and bubble density) under a variety of operational conditions, an increase in the ability to disperse air in the flotation cell, a reduction in the rate of coalescence of individual bubbles in the cell, and a decrease in the rate of bubble rise to the pulp surface. It will not be the intention of this paper to exhaustively list all frother accomplishments related to research, development and use in coal flotation. Rather, the authors will attempt, based on their experience, to present a technology overview of the role of frothers in the chemistry of flotation as successfully practiced at the industrial level. A very useful framework for flotation is the concept suggested by Klimpel (1984, 1984a) of viewing flotation as a system consisting of three major components: chemistry, equipment and operating variables. This is illustrated in Figure 1. A major point of this figure is that flotation is a highly interactive system with many technical and economic combinations of component settings available to operators to achieve a given goal. Sometimes simultaneously changing various component settings will reinforce a particular attribute. Also various component settings can cancel or counteract each other if changes are not chosen wisely. Thus, coal flotation must be viewed by the plant operator as a complex system with the economic/technical optimization of a given flotation process being directly related to the operator's knowledge of how his particular flotation system behaves as a function of the major control variables he has at his disposal. Typical important component settings in coal flotation are collector dosage, frother type and dosage, air flow rate, feed rate, particle size, pulp density, and the surface oxidation conditions of the coal. Because of the above complexities, it is sometimes assumed (especially in laboratory studies) that the choice of frother type and dosage is not as crucial as some of the other factors such as collector type and dosage (e.g., Leja, 1982). Detailed plant level testing of all of the factors of Figure 1 performed in recent years, Klimpel (1980, 1980a, 1982, 1984, 1984a), Hansen (19861, Klimpel, Hansen and Meyer (1982), simply do not support this assumption. The above testing has shown rather clearly that certain frother chemical structures consistently exhibit predictable performance characteristics in both mineral and coal flotation. In addition, significant changes in both rate and recovery can be expected. In the remaining sections of this paper, the following topics will be covered: requirements of chemicals capable of use as frothers, types of chemicals used commercially as frothers, some general performance characteristics of existing commercial frothers, and recent work in developing new frother chemical structures.

Jan 1, 1986

By Manfred Hager

The ability of conveyor belts to transport large mass throughputs economically over previously unprepared ground has resulted in this system achieving great and extensive use. A significant component in this development is the conveyor belt itself. The development of high strength steel cord conveyor belts involves the optimising of splice design, the use of excellent rubber material especially in the splice, good craftsmanship during splice production and modern field vulcanisation equipment. The durability of a splice for belts of class St 3000 up to St 790 is expressed by the fatigue strength under dynamic stress. The results obtained with a test method and equipment developed by the University of Hannover indicate the present state of the art in this field. Belts of high nominal strength used on inclined long distance conveyors reach splice fatigue strengths of about 3000 N/mm. The design should only exploit approximately 50% of the fatigue strength determined in the tests as maximum operating stress. IMPORTANCE OF CONVEYOR BELT SYSTEMS The use of conveyor belt technology has expanded considerably over the last decades in the bulk goods transportation sector. Because of the favourable transport costs and the technology's adaptability to specific topographies, belt conveyors frequently represent the most economic solution. When flows of goods are large, as in German lignite mining operations, where masses of up to 40,000 t/h are given, it is the only technically feasible alternative. The use of belt conveyor systems allows the sensible distribution of the waste and the production of the coal (Hager, 1981). But this transportation alternative is also suitable when the masses involved are smaller. Whereas conveyor belts were only suitable for loose bulk goods up until a few years ago, today they are frequently also used for hard rock open cast mines, e.g. in the production of copper ore. Because of the favourable transport costs it is frequently still economic to use mobile crushers with a throughput of up to 10,000 t h 6 which breakdown the large blocks produced y explosives into transportable grain sizes. In many operations, trucks are only used to provide flexibility between the excavator and the crusher, i.e. over short routes, and the long, frequently steep, transport paths to the processing plant or to the spreader are undertaken by belt conveyors (Einenkel et al., 1992). The advantages of belt conveyors, i.e. to cope with large mass flows over inclines and down slopes of up to 1:4 have resulted in this system gaining very extensive use throughout the world. The conveyors can be installed over for the most part unprepared ground with suitable vertical radii, adapted to the locally available belt material. In the process individual conveyor belts with lengths of up to 15 km, and in the underground sector with lift heights of up to 1 km, have been built and operated. This great variety of application possibilities is complemented by a limited ability to pass through horizontal curves. This can be achieved with the help of design measures on the conveyor whilst bearing in mind the characteristics of the belt. STRESSING OF CONVEYOR BELTS IN DIFFERENT TYPES OF PLANT The advantage of conveyor belts compared with other systems is also to be found in the large available service-time window, i.e. the excellent reliability and the low costs of energy consumption and maintenance. It is in particular the reliability of a belt transport system which depends to a great degree on its main component, that is the belt - in all its different types and variations. For this reason, the belt is also given the greatest attention in the development of the individual components, because it is the belt which must be designed optimally with regard to various factors. The specific properties of the conveyor belt then assert a considerable influence on the design and sizing of most other components in a

Jan 1, 1993

By Donald C. Moore

Introduction Asamera Minerals (US) Inc. and its joint venture partner, Breakwater Resources Ltd., discovered ore grade gold mineralization on their 20 km2 (5000 acre) Wenatchee, WA land position in February 1983. Due to the high grade nature of the discovery ore and the known reserves of ore in the "B Reef' and "B West" zones previously outlined by other companies, a decision was made to construct a mine/mill operation near the known ore occurrences. Further drilling in the discovery area quickly expanded known gold occurrences to more than 3.6 Mt (4 million st) with tentative in-place ore grade of 7 g/t (0.25 oz per st) and minor silver values. Based on existing knowledge of the ore body and the rapidly increasing ore reserve, a decision to build a 1.8-kt/d (2000-stpd) mine and mill complex was made in the second quarter of 1983. A schedule was devised to begin immediate mine development, shaft sinking, environmental and land use permitting, and mill and tailings dam construction (Fig. 1). Meeting the scheduled startup date, April 1, 1985, required a fast track schedule in all areas. To this end, Asamera purchased the Oracle Ridge Partners concentrator. This was an assemblage of new equipment designed for use as a copper concentrator in southern Arizona. The purchase contained all of the major mineral dressing equipment - crushers, screens, rod and ball mills, etc. and an engineering package. It did not include most of the other required items, such as buildings, conveyors, pipelines, tanks, and pumps. At the same time, core samples were sent to two independent process development laboratories for initial flowsheet development. Due to the refractory nature of the carbonaceous ore, cyanide leaching was not feasible. Flotation was selected as the concentration process. Further testing showed that autoclaving of the flotation concentrate followed by cyanidation would result in overall recovery of about 85% gold. A mine manager was hired to begin assembling an operations staff, hire an environmental consulting firm, and begin mine development. Environmental and land use concerns were major obstacles due to the mine's close proximity to a city of 20,000 people. These concerns had to be rapidly defined so as to mitigate any adverse impacts from and mining processing operations. Baseline data dealing with weather, air and water quality, and sound were measured before start of mine construction. Concentrator and flowsheet development remained static until October 1983 while definition drilling and mine development proceeded. In late October, a process engineer was hired to coordinate development of a process flowsheet, purchase the remainder of the concentrator equipment, prepare a concentrator construction contract, finalize concentrator detail engineering, and combine environmental and process requirements with a tailings dam design. Process development There were only 17 months remaining to mill start up from the hiring date of the process engineer. Therefore, the process flowsheet had to be finalized rapidly. To accomplish this, samples of drill core from the highest grade (and therefore potentially the most commercial) ore zones were sent to an outside metallurgical laboratory to confirm beneficiation tests on the flotation process. Test results again showed that flotation would provide about an 86% gold recovery. Therefore, all further testing was concentrated on flotation and autoclave/cyanidation of flotation concentrates. Focusing on a well known process such as flotation was important in accomplishing the rapid design and construction of the concentrator. If, during these next phases, we were continually changing design concepts, layout, and process flow, the mill startup would have been delayed many months. Once a process flowsheet is selected the process engineer must obtain the process criteria needed to design the beneficiation system. For example, it was known in early December that the Oracle Ridge rod and ball mills were too small to grind 1.8 kt/d (2000 stpd) of Wenatchee ore. A decision had to be made to purchase a large, used ball mill and convert the Oracle Ridge ball mill to a rod mill. The process engineer must be cognizant of the process criteria needed to size and select equipment. If not, the process engineer must use the professional services of the equipment manufacturing companies to review the requirements that the equipment is asked to perform. For the Wenatchee system, this resulted in the adaptation of a ball mill to a rod mill with a weight limit of grinding rods to protect the mill bearings and drive trains. When a decision is required, the process engineer has to present the facts and options in a manner that allows a rapid decision. This information must include costs, equipment availability, and effect on the construction schedule. At the Cannon mine, there were process development details that resulted in decisions similar to the ball mill purchase. These included an increased flotation residence time from eight to 25 minutes, an increased thickener area requirement, a high pressure tailings pumping system, and area constraints in plant layout. All of these decisions had to be timely and required assistance from manufacturers' service engineers, and knowledge of the alternate costs and effects on construction completion. Equipment procurement It was decided in early 1983 to build the ore milling facility with Oracle Ridge equipment, augmenting it with used equipment

Jan 2, 1989

By Thomas D. Henderson, Wayne R. Hopkins, Arthur Lynch

Brought on stream in 1975, the 10,000-tpd copper oxide processing plant of Anamax Mining Co. at Twin Buttes, Ariz., is now at design capacity. Based on acid agitation leaching, solvent ion exchange, and electrowinning, the Anamax plant bridges an ore to metal gap by producing mine site copper cathodes from lime-bearing oxidized feed. Until recently, such material was considered submarginal because copper recovery was uneconomic using conventional methods. The new Twin Buttes oxide plant, the largest of its kind in North America, is one stage in the expansion of Anamax, a joint venture owned equally by The Anaconda Co. and Amax Inc. Other phases include an accelerated stripping program in the pit, an expansion of the sulfide concentrator from 32,000 to 40,000 tpd, and construction of a 57 MW diesel generating station for the oxide plant. Once the basic oxide flowsheet had been selected, Anamax awarded a contract to a joint venture of Arthur G. McKee and Co.' (Western Knapp Engineering Div.) and Davy Powergas' (a member of the Davy International group). The contract covered process design, engi¬neering, procurement, and construction. The property, identified by Banner Mining Co. in the 1950s and developed by Anaconda, was put into production in 1965 as an open pit mine equipped with a 32,000-tpd sulfide flotation concentrator. The Twin Buttes mine has since become one of the largest earth¬moving operations in the world, excavating some 10 tons of waste per ton of ore. At the original sulfide mill feed rate, the mining fleet moved 375,000 tpd of material (225,000 tpd alluvium and 150,000 tpd of rock). The most significant copper mineral in the ore body is chalcopyrite, which is scattered in pockets of weak mineralization. Cap oxidation has progressed to depths of several hundred feet, principally as chryso¬colla. Present estimates place minable reserves at 347 million tons of sulfide ore grading 0.6% copper and 55 million tons of oxide ore grading 0.9% recoverable copper from 1.25% total copper. The nonre¬coverable fraction is composed of native copper and insoluble copper oxides and sulfides. The oxide ore has been partially stripped, and 25 million tons having a copper content greater than 0.6% have been stockpiled. The stripping of oxide ore to expose sulfides will continue for several years. The copper oxide studies launched by Banner Mining in the 1950s were continued by Anaconda. Process routes investigated included: acid leaching; cyanide leaching; segregation roasting; chlorination, oxidation, and reduction; flotation in combination with leaching; cal¬cite flotation plus leaching; copper oxide flotation; and sulfidization flotation. By late 1968 test work favored a low temperature roast followed by an ammonia-ammonium carbonate leach of the ore. During this period, acid leaching had been discounted because of excessive con¬sumption of expensive acid by the high limestone ore matrix. Restric¬tive SO2 emission controls in 1971 suggested an increase in acid production and availability at a lower cost. Research then switched to acid leaching, resulting in development of the present flowsheet. Mechanically agitated leaching in a cascaded series of tanks was selected in preference to vat leaching after piloting a stirred-tank system. Preliminary testing of vat leaching indicated that inter-vat solution clarification problems could be caused by gas evolution from leaching of high limestone ore. Lack of research time for resolution of this problem, plus success of the pilot agitation system, dictated the final flowsheet selection. In 1973 Anamax decided to proceed with agitation acid leaching to produce liquor for a solvent ion exchange (SIX) and electrowinning (EW) installation. At that time, such systems were proving their eco¬nomic viability, and they allowed direct recovery of high quality cath odes while bypassing the smelting and refining of lower grade cement copper precipitates. The Anamax oxide plant processes a grade of about 1% recovera¬ble copper from stockpiled sources and new mine production. Unit operations include size reduction, agitation leaching, countercurrent decantation washing of residue, pH adjustment of solution, clarifica¬tion, solvent ion exchange, and electrowinning of copper, with a mini¬mum of environmental impact. The feed to the primary crushers is composed of: 1) -10-in. stockpiled oxide ore, which has passed through the mine crusher and has been grade-segregated to allow blending for a consistent feed material. 2) -4-ft ROM oxide ore stockpiled in the early mining operation. 3) -10-in. ore currently being mined. Ore is delivered in 100-ton mine trucks to a dump pocket, where an apron feeder transfers the ore to an inclined 6-in. fixed grizzly. The grizzly oversize passes through a 48 x 60-in. primary jaw crusher, where it is reduced to -6 in. The fixed grizzly undersize and the crusher product are conveyed to a 4-in. vibrating grizzly. Oversize from this grizzly reports to a 7-ft Standard cone crusher, where it is reduced to -1 in. The crusher product and the vibrating grizzly undersize are conveyed to two 8 x 20-ft double-deck vibrating screens that operate in closed circuit with two 7-ft short head cone crushers. Screen undersize (-1/2 in.) is delivered to a covered conical fine ore stockpile with a designed live capacity of 15,000 st. Beneath the stockpile are ten Pioche belt feeders feeding two conveyor belts that supply the milling circuits. Each group of five feeders has two variable speed and three fixed speed drives. Grinding to 95% -48 mesh is done in two parallel lines, each consisting of an 111/1 x 18S -ft rod mill and a 12' x 30-ft ball mill in open circuit. Water is added to maintain a pulp density of 72% solids in the mills. Further water additions are made to the ball mill discharge sumps to create a 60% solids slurry for transfer to the leaching reactors. The nominal grinding rate is 440 tph solids. A splitter box on the product from the mills separates the incoming slurry into two streams, one for the leaching reactors and the other for the pH adjustment reactors. This split is determined by the lime content of the feed material and the pH of solution entering the pH adjustment reactors. Five 30-ft diam by 31-ft high mechanically agitated, rubber-lined leach tanks are arranged in cascade, with gravity transfer from one to another by enclosed launders. Concentrated sulfuric acid (93.2%) from storage is added to the pulp principally at the first leach tank, and the reacted slurry leaving the cascade after a 5-hr residence time at 50% solids is pumped to countercurrent decantation (CCD) wash thickeners.' At this point, most of the soluble copper has been leached from the ore. Up to 250 lb of sulfuric acid is consumed per ton of ore, the acid being supplied in tank cars from a nearby smelter. Leached slurry at 50% solids is washed countercurrently with return raffinate from the SIX plant in a series of four Doff-Oliver center caisson, 400-ft diam thickeners. Solids advance from thickeners No. I through No. 4 to the tailings dam. SIX raffinate advances from thickener No. 4 to thickener No. l, where a pregnant solution overflows to pH adjustment. To optimize solvent extraction, the pH of the pregnant solution from CCD thickener No. I is adjusted from 1.5 to 2.5 by adding unreacted ground slurry (split from the grinding circuit) to the preg¬nant overflow of No. I CCD thickener. The resulting slurry is retained for 45 min in three mechanically agitated leach tanks. CO, generated during leach and pH adjustment is ducted to a demister, where entrained acid droplets are removed. The cleaned CO, and the No. 4 CCD underflow are the only effluents from the plant. The underflow of No. 4 thickener mixes with tailings from the sulfide concentrator and is neutralized. Slurry exits the pH adjustment vessels at about 10% solids and

Jan 1, 1985

By E. F. Klein, R. L. Eacret

Introduction In contrast to the general type of crushing mechanism in which the crushing surfaces alternately approach and withdraw from each other, described earlier in this section, and continuous-pressure break¬ers such as rolls and roller mills that will be described in a later chapter, impact crushers load by striking pieces of rock while in free fall and hurling them at high speed against stationary surfaces. Because the impact crusher depends for its effectiveness upon high velocity, wear is greater than in the slower-moving jaw and cone¬type wear parts, and for this reason its use is strictly limited to rela¬tively soft, friable, and sticky rocks that are characteristic of many nonmetallic mineral deposits. A few of these are phosphates, lime¬stone, barite, clay, asbestos rock, and coal. However, several examples of their use on soft lead-zinc and precious metal ores have been known. Although the use of impact crushers is expanding today into the range of ores containing as much as 15-18% silica, Taggart16 set the practical limit at 5%, and in the 1940s and 1950s several installa¬tions in the US and western Europe exceeded the limits of economical maintenance and were quickly superseded by the slower-moving, con¬ventional crushers. A rock that tends to be plastic or bouncy in a jaw or gyratory crusher when the force is applied slowly to reach ultimate strength, may become brittle with rapid loading and thus increase the effective¬ness of the applied forces. For this reason it is to be expected that as the quality of hammers, grates, plates, and cages improves with advances in steel technology, the use of high-speed crushers of this kind will increase. Terminology Modern usage differentiates between the impactor and the hammer mill, the former relying primarily on the impact of hammers (fixed or free-swinging) and secondarily upon pieces striking one another or steel surfaces; the hammer mill relies on both the centrifugal impact force of free-swinging hammers and the attrition and shear action between these hammers and well-placed grates suspended at the bot¬tom just below the hammer circle. The hammer mill, because of its grate discharge, restricts discharge of oversize rock to the grate open¬ing, while at the same time providing a trap for removal of tramp iron or other uncrushables. The impactor discharges free, so generally works with a screen to control product sizes. The question of terminology, impactor vs, hammer mill, creates difficulties because the similarities appear to outweigh the differences by far; if one were to list the similarities in order of importance and then the differences, he would be forced to conclude that they would best be dealt with as a single kind of crusher. Taggart16 gave it four names and added "as it is variously known," but it must be remembered that in 1945 the machine was nearly exclusively of the flailing-hammer type, while today the fixed-hammer rotor is also com¬mon. In this chapter the terms impactor and hammer mill will be used where they seem to apply. It is perhaps unfortunate that this terminology is being confused with rock breaking at the mine, usually with hand-held tools, e.g., the article "High-Energy Impact Rockbreaking" by Grantmyre and Hawkes, CIM Bulletin, August 1975. General Description Impact breakers, impact crushers, and hammer mills accomplish material breaking and reduction primarily through impact action of the material with fixed or free-swinging hammers revolving about a central rotor. The material to be crushed enters through an opening at the top or top side known as the "feed opening" or "hopper opening" and falls into the path of rotation (hammer circle) of the hammers. Initial breakage is accomplished in midair by collision of the dropping feed material with high-speed hammers. The second stage of breakage occurs when the pieces hit plates or breaker bars which line the crusher boxlike frame. Hammer mills rely further on a shearing and attrition action between free-swinging hammers and grid bars or grates at the crusher bottom which restrict discharge of oversize material until it is broken sufficiently to pass through the grid opening. The term hammer is used in reference to the piece which strikes the material, whether it is fixed on the rotor or free-swinging. It

Jan 1, 1985

By A. H. Leuschner, P. J. Kruger, J. Kruger

INTRODUCTION This paper deals with some practical aspects of the use and interpretation of dosimetric methods for assessing the exposure of workers to natural uranium. Consideration must be given not only to the scientific value of the method of dosimetry, but also to the practicality of the method in the particular working environment. It is generally accepted (IAEA, 1976) that uranium does not present a hazard during the mining and extraction processes up to the final concentration stage, that is precipitation as ADU or calcining to the oxide. Further processing (e.g. conversion to oxides, fluorides, etc.) is generally associated with the nuclear industry. It is customary for uranium processing facilities to be managed according to industrial norms rather than according to those applied for the purpose of radiation protection. This is mainly because of the large quantity of material to be processed and the fact that uranium is considered to be of a low radiotoxicity. Although ventilation is used to some extent, provision is not made for the same level of protection as would be required in a radiochemical facility. It is typical in such a plant to find that the process itself may be enclosed and ventilated, but the ventilation is inadequate to cope with an accidental release of material. The working environment is poorly ventilated, sometimes only by natural ventilation, and one frequently finds areas with high airborne dust levels, designated as mandatory respirator areas. Protection of personnel is dependent on personal protective equipment. Radiation protection is not of prime concern and personnel are not specifically trained in this subject. It is in this type of environment that acceptable monitoring procedures and personnel dosimetry must be established. The effectiveness and practical application of the dosimetry must also be judged against this working environment. DOSIMETRIC TECHNIQUES The methods available for assessing personnel exposure, whether it be in terms of chemical or radiological hazard, include urinalysis, faecal analysis and [in vivo] monitoring. Faecal analysis does not present itself easily as a method for routine use. Techniques for [in vivo] monitoring have been developed recently and as expensive instrumentation is required, it is not generally available. That leaves urinalysis as a dosimetric technique for routine use, and it is probably for this reason that urinalysis is still widely used. Guidelines for the interpretation of urinalysis results, as originally provided by Neuman (Neuman, 1950), are still used in practice, even though it has become clear that this method has severe limitations as regards the assessment of the dose as a result of the intake of class W or class Y compounds (Alexander, 1974). Typically a level of 300 µg [U/C] urine can serve as an indicator of an acute exposure (above which chemical damage to the kidneys may occur) and a level of 100 µg [U/C ] urine can serve as an indicator for an investigation. Such levels are determined and are used in conjunction with several factors, such as mode of intake (ingestion or inhalation), solubility of compound (D, W or Y), means of intake (acute or chronic), environmental monitoring results and frequency of urine sampling. Since the ICRP concept for internal radiation limitation changed from that of the critical organ, i.e. the single organ of greatest significance under the circumstances, to that of effective dose equivalent, i.e. account being taken of the total risk due to the exposure of all tissue, the maximum permissible organ burden (MPOB) was replaced as secondary dose limit by the annual limit on intake (ALI) (ICRP, 1977). The ALI values are calculated from the committed dose equivalents of the various organs, and are used to determine organ burdens, which are then used to interpret the dosimetric results. Using the ICRP model (ICRP, 1979), a comprehensive calculation was made by Johnson (Johnson, 1980), giving organ burdens and excretion data in terms of intake for acute and chronic exposures for different categories of uranium compounds. The question arises as to what extent the direct measurement of intake can be utilised as a method of dosimetry. This will require a method of personal intake measurement for each individual worker, with both ingestion and inhalation being taken into account. A measurement of

Jan 1, 1981

By Guangming Zhao, Nianjie Ma, Demao Guo, Denghong Chen, Yingming Li

Based on the mechanical essence that large-scale plastic failure zone appears in all or part of surrounding rock in soft rock roadway, the numerical simulation method is used to study the rectangular roadway in layered rock strata. It is clarified necessary conditions must be met for soft rock: firstly, the strength condition is that the maximum confining pressure is greater than the uniaxial compressive strength of rock strata. Secondly, the stress environment condition is that the ratio of maximum confining pressure to minimum confining pressure is greater than 3. Thirdly, the angle condition is that The direction of principal stress action enables the plastic zone of weak rock layers to fully develop. At the same time, the quantitative description method of soft rock is given, and the soft rock roadway is redefined. Soft rock roadway refers to the roadway that meets the strength conditions, stress environment conditions, and rock structure angle conditions under certain surrounding rock conditions and in-situ stress environment conditions. After the excavation of the roadway, a large-scale plastic failure can be formed, that is, a butterfly-shaped plastic zone is formed, and the conventional support cannot be adapted. It is difficult to support in engineering. It provides a theoretical basis and engineering analysis method for the identification of soft rock roadway, and the research results have engineering value Soft rock tunnel engineering in coal mines constitutes a vital aspect of soft rock engineering. This field broadly encompasses rock engi- neering concerning large plastic deformations, e.g., soft rock slope engineering and soft rock tunnel engineering. The intricate geological conditions encountered in soft rock tunnel engineering present a significant challenge to support, which has harmed coal production in China. China leads global raw coal production with the annual output of 4.6 billion tons. Annual tunnel excavation supporting this production spans approximately 11,000 km, with over 10% of these tunnels classified as soft rock formations. Soft rock is commonly associated with soft rock tunnels due to their prevalence in engineering projects. However, reaching a consensus on the definition of soft rock has long been an enduring challenge for scholars and engineers. Numerous definitions have been proposed, includ- ing descriptive, index, and engineering definitions. For instance, the International Society for Rock Mechanics defines soft rock based on its uniaxial compressive strength σ ranging from 0.5 to 25 MPa. China's Engineering Rock Body Standards, established in 1994 (GB 50218-94), take a qualitative and quantitative approach to classifying rocks. Rocks are categorized as hard or soft based on criteria such as hammering sound, fragmentation, water immersion effects, and weath- ering degree. Additionally, the integrity of rock bodies is assessed across five categories intact, relatively intact, soft fractured, fractured, and extremely fractured. This classification considers factors like the number and spacing of structural planes, their combination, and the types of structures. Descriptive and index-based definitions fall under the category of geological soft rocks, providing a comprehensive geological perspective on the surface features or strength characteristics. However, these definitions have limitations in engineering practice, which leads to contradic- tions. For instance, rocks with uniaxial compressive strength less than 25 MPa may not exhibit soft rock characteristics if the tunnel is shal- low with low horizontal stress levels. Conversely, rocks with compressive strength exceeding 25 MPa at sufficient depth and high horizontal stress may exhibit soft rock characteristics. Definitions originating from engineering practice have emerged after realizing the inadequacy of discussing soft rocks without considering engineering. For instance, Dong's loose circle theory defines soft rocks as rocks with a loose circle thickness exceeding 1.5 m, which chal- lenges conventional supports. This intuitive definition, widely accepted by engineering professionals, emphasizes the difficulty in supporting tunnels due to extensive damage. However, various tunnel damage poses a challenge in relying solely on the loose circle thickness of tunnels for determining soft rocks. He introduced the concept of engineering soft rocks, which are defined as rock formations exhibiting significant plastic deformations under applied engineering force. Two fundamental mechanical properties of soft rocks are identified the critical softening load and critical soft- ening depth. Rocks below the critical softening load threshold are categorized as hard rocks, while those exceeding it exhibit substantial

Jun 25, 2024

By William Cass, Ellen Caftel Turcotte

The Elliot Lake Centre was established in 1965 as an independent Centre for Continuing Education offering a variety of adult education services. In 1976 all retraining and community college type activities were transferred to the Sault College of Applied Arts and Technology which had opened a branch campus in Elliot Lake. Upon the completion of this transfer, the Elliot Lake Centre was then contracted in 1977 by the Canada Centre for Mineral and Energy Technology to study the feasibility of establishing a research and information institute dealing with various health and safety concerns of the uranium worker. By late 1980, a federal charter was granted for the establishment of the Canadian Institute for Radiation Safety, made possible through the cooperative efforts of the Governments of Canada and Ontario, the Ontario uranium mining and refining companies, and the Elliot Lake Centre. On 1 April 1981 the Institute officially began operations. PROGRAM DEVELOPMENT AND COORDINATION In May 1978 the Elliot Lake Centre initiated a voluntary sputum cytology program for uranium workers. This program was financed by the Ontario uranium mining companies: Agnew Lake Mines Limited near Sudbury; Madawaska Mines Limited in Bancroft, and Denison Mines Limited and Rio Algom Limited in Elliot Lake. In 1981 additional funding was granted by Eldorado Nuclear Limited for a small group of employees at the company's uranium refinery in Port Hope, Ontario. Initially in 1978, program participants were accepted according to length of employment, radiation exposure levels (calculated in working level months) and/or smoking history. These restrictions were abolished within a few months and all uranium workers were encouraged to participate. From 1978 to 1979 the sputum cytology program was medically directed by personnel from the Montreal Cancer Institute. Until 1979 all data maintenance for the program was performed manually, however the effort required to sustain the system increased significantly as the number of participants grew. In September 1979 the Division of Thoracic Surgery at the Toronto General Hospital, under the direction of Dr. F.G. Pearson, submitted a proposal to the Elliot Lake Centre to develop a data processing system to help coordinate the surveillance program. In October 1979, Dr. Pearson became the new Project Director. Director of the Cytology Program is Dr. D.W. Thompson of the University of Toronto who is responsible for all cytologic analyses. Developmental work to establish the new surveillance system was initiated in January 1980 and became fully operational at the beginning of March 1981. The following sections describe the Elliot Lake operations of the sputum cytology program now conducted by the Canadian Institute for Radiation Safety. PROGRAM PARTICIPATION As of 1 September 1981 the sputum cytology program has 518 participants enrolled. Eligibility is restricted only to a willingness to participate and the ability of an individual to produce adequate sputum samples. Each participant registers for the program at the Institute and is given personal instruction on methods of sputum collection. The majority of the participants in this program are employed by the uranium mining industry; some work for firms which are contracted by the larger companies. Many participants are either retired from the work force or receive disability pensions or workmen's compensation benefits. Other participants have been employed in the uranium mining industry and have since gone on to other employment; some work for various federal and provincial governmental agencies concerned with the mining industry. Some participants are referred to the program by their family physicians because of smoking histories and/or various respiratory conditions. A small number of these individuals have never been involved in the mining industry at all. A chart describing program participation by employment is found in Appendix 1. PROGRAM REGISTRATION In the Elliot Lake area, persons interested in registering in the program are required to go directly to the Sputum Cytology office located at the Canadian Institute for Radiation Safety. Each individual is interviewed personally by the Program Officer. For the employees of Agnew Lake Mines, Madawaska Mines and the Port Hope refinery of Eldorado Nuclear Limited, registration is possible during the Program Officer's semi-annual on-sitevisits. Participants are assigned, in chronological order, "program numbers" for easier identification. File labels are color-coded according to place of residence or, in some instances, place of employment. Each file is stamped with the date each sample bottle is distributed. A file contains: all original test results, a completed registration

Jan 1, 1981

By Mark E. Smith, Gerald J. Gierszewski

Introduction The purpose of this paper is to present a procedure used by the authors for evaluating and designing soil liner systems. This method is particularly valuable in evaluating various treatment schemes for chemically amended soil liners. A tabulation of laboratory test results on various soil types are presented to quantify the effectiveness of certain treatments. A typical liner design program includes developing and proving soil borrow sources, designing the cross-section of the liner system, developing construction specifications, and providing construction services to ensure the intended product is achieved. Material Source Development The first step in designing a soil liner is to identify and evaluate suitable borrow sources within an economical haulage range. This is best done in a two step approach: a reconnaissance level investigation to identify target areas and a detailed evaluation of those targets. Reconnaissance: The goal of the preliminary investigation is to locate potential borrow sources for liner quality soils. This includes all natural materials which can be compacted, chemically treated, or otherwise amended to yield an installed permeability at or below some target value. This requires utilization of all available data sources: Soil Conservation Service, BLM, aerial photos, USGS geologic maps, and project geologist records. The goal at this stage is to locate shallow deposits of favorable soil types. The Unified Soil Classification System provides an excellent first pass grouping. Clays, clayey sands and silts are the most favorable soil types, although silty sands and occasionally clayey gravels can make excellent liners, and are often amenable to chemical modification. The lowest permeabilities are generally achieved with CH, CL and MH soils. Once preliminary targets have been identified using visual examination, laboratory classification tests should be performed to further refine the selection. Testing at this stage should include gradation, plasticity and hydrometer analyses. Additionally, "preg-rob" testing should be done as early as practical. Preg-rob is a phenomenon where gold or silver ions in solution associate with the clay, or other, minerals. When this occurs, a portion of the gold or silver leached from the ore is actually tied-up by the clay and thus a reduces recovery. Testing for this consists of agitating a small sample of the soil in a solution containing dissolved gold or silver, preferably of similar chemical make-up as the solution which will contact the actual liner. The solution and soil are assayed before and after agitation to determine loss to the clay. A reliable estimate of the hydraulic conductivity, commonly referred to as permeability, can be developed from the D10 value by the use of Hezen's formula: K = 100 (D1012 This relationship is limited to soils where the finer particles do not move due to the force of flowing water (i.e.: "hydrodynamic stabilitym)(1). Additionally, the effect of platty particles on permeability is not as predictable as the effect of equidimensional particles. D10 is the sieve opening size at which 10% of the material is finer. Plasticity is also important from several standpoints. Constructability is directly related to plasticity. Very plastic clays and non-plastic silts both tend to be difficult soils, while medium plastic clays and clayey sands are generally very desirable. Post construction performance is also related to plasticity (e.g. swelling, shrinkage cracking, frost heave, etc.). Additionally, low plasticity silts and silty sands generally do not respond well to chemical amendment. Source Development: The result of the reconnaissance evaluation should be an estimate of the relative probability of developing a suitable borrow source within an economical haul distance. Of course, "economical distance" depends on the degree of handling and treatment the borrow material requires, as well as the cost of synthetic alternatives. The purpose of the detailed investigation is to prove out quantity and quality of material sources, and determine design parameters such as degree of compaction, mixing, treatment and thickness of liner. The emphasis of the testing program should be permeability and strength. Strength becomes increasingly important as the slope of the liner and the height of the heap increase. Permeability testing should evaluate the effects of compaction, water content, mixing and chemical treatment where appropriate. The effects compaction and water content during compaction have on

Jan 1, 1987

By G. F. Raymond

Discussion by G.F. Raymond Knudsen's study presents two curious conclusions: • The kriging of blasthole assays can systematically overstate mill head grades by as much as 21% as a result of unbiased sample variance. • No estimation method is able to reduce this bias to a small margin. The study is based on a simulation using real data and what are presumed to be actual variogram parameters from a real deposit. Although I have no doubt that the first conclusion (21% overestimation) is correct for the author's simulation, I do not believe this represents a realistic mining situation. Over the past 15 years I have done extensive comparisons between exploration drill-hole assays, blasthole assays and mill head grades on seven major open-pit mines, including some very erratic gold deposits. Commonly, nugget effects on blasthole variograms were 10% - 20% higher than on exploration variograms. And in one extreme case, the difference was 50%. Even in the extreme case, ordinary kriging on blastholes agreed well with the mill head grade over the long term. In Knudsen's simulation, the blasthole nugget effect is assumed to be 200% higher than exploration data's. He supports this by variogram plots from each. My guess is that the apparent, large difference between these variograms results from a failure to account for the proportional effect (blasthole assays are likely from a higher-grade area). A simple check would be a comparison of the variance of exploration samples nearest blastholes. As for a nearly conditionally-unbiased estimator of a large random error, the arithmetic mean of all of the data certainly qualifies, provided there is an even data spacing. As a corollary, so does simple kriging, which would include, in this case, a large weighting to the arithmetic mean. Similarly, using a large number of samples with ordinary kriging or indicator kriging would significantly reduce the bias in the case of a large nugget effect for the variogram.[ ] Reply by H. Peter Knudsen Raymond questions two conclusions in the paper. First, he wonders whether a 21 % overestimation represents a realistic mining situation. I agree that 21 % is high, but overestimation in the range of 10% to 15% is certainly common in my experience. Furthermore, several years ago I consulted on a gold mine that was experiencing a 45% overestimation due, predominantly, to poor blasthole samples: In further questioning of the 21 % value, Raymond wonders whether the nugget effect of the blastholes is really so much larger than the nugget values of the exploration data. It is consistently larger throughout the deposit. In my experience with six Nevada gold mines, the high nugget value is not unusual. In fact, for some reason, nugget values for blasthole samples are typically about 0.0005 (opt squared). I am of the opinion that this high nugget effect observed at many gold mines is predominantly due to the inherent inadequacies of the blasthole sample and subsequent sample preparation. The second conclusion Raymond questions is the inability of the estimators tested to reduce the conditional bias. In fact, the conditional bias is extreme with the polygon estimator and greatly reduced by ordinary kriging. However, it was not eliminated. Raymond suggests using simple kriging, or perhaps a larger number of samples, to reduce the conditional bias. The technique of simple kriging may be less affected by the random errors in the data, but I did not test the technique. Using a larger amount of data presupposes that too few samples were used initially. In ordinary kriging and indicator kriging, the screen effect comes into play and ensures that samples beyond the second screen are given zero weight. Hence, increasing the sample size does not change the estimates nor the conditional bias. The main point of my paper is that the random unbiased errors (a fact of life in blasthole samples) cause a conditional bias in our estimates. The mechanics by which the conditional bias is introduced are nicely explained by Springett. My paper simply shows that the bias is also present when working with linear estimators, such as ordinary kriging, and even with nonlinear estimators, such as indicator kriging.[ ]

Jan 1, 1993

By Chester A. Rowland, David M. Kjos

INTRODUCTION Mineral ore comminution is generally a feed preparation step for subsequent processing stages. Grinding, the fine product phase of comminution, requires a large capital investment and frequently is the area of maximum usage of power and wear resistant materials. Grinding is most frequently done in rotating drums utilizing loose grinding media, lifted by the rotation of the drum, to break the ores in various combinations of impact, attrition and abrasion to produce the specified product. Grinding media can be the ore itself (autogenous grinding - primary and secondary), natural or manufactured nonmetallic media (pebble milling) or manufactured metallic media - steel rods, steel or iron balls, or a combination of autogenous media and steel balls (SAG milling). This chapter covers rod and ball mills utilizing manufactured metallic grinding media. MILL DESIGN The interior surface of rod and ball mills exposed to pulp and/or grinding media are protected from wear and corrosion by rubber, metallic or a combination of rubber and metallic wear resistant materials. Rod and ball mills essentially draw constant power, thus are well suited for use of synchronous motors with power factor correction capabilities as drive motors. A net of approximately 120 to 130 percent of running torque is required to cascade the charge in these mills. The pull-in torque is about 130 to 140 percent with the pullout torque to keep the motor in-step (in-phase) generally in excess of 150 percent. When rod and ball mill are started across-the-line the starting and pull-in torques result in inrush currents exceeding 600 percent which possibly result in high voltage drops. To deliver 130 percent starting torque to the mill the motor design must take into account the maximum anticipated voltage drop. Motor torque decreases as the decimal fraction of the voltage available squared. E.g., a motor rated 160% starting torque with a 10% system voltage drop will deliver 160% x or 129.6% torque to its output shaft When it is not possible or practical to start a fully loaded synchronous motor across-the-line it is possible to utilize the motor's pull- out torque to start the mill. By using a clutch, normally an air clutch. between the motor and the mill, the motor is brought up to synchronous speed before the clutch is energized. If the motor has an adequate amount (175 or greater) of pull-out torque the pull-out torque starts the mill without major disruptions on the electrical system. Since the energy release at initial cascade of the mill charge is an inverse function of acceleration time, a minimum acceleration time of 6 to 10 seconds or more is recommended to prevent damage to the mill or the mill foundation. Economics at the time of plant design and mill purchase determine the drive to be used. The simpliest drive is the low speed synchronous motor with speeds in the range of 150 to 250 RPM connected to the mill pinionshaft by either an air clutch or flexible coupling. Using a speed reducer between the motor and pinionshaft permits using motors having speeds in the range of 600 to 1000 RPM. In this speed range, if power factor correction is not required induction motors can be used; squirrel cage where there is no restriction on inrush current; slip ring where a slow start and low inrush current is required. Air clutches can also be used to ease starting problems with squirrel cage motors. In some areas of the world induction motors and starters are less expensive than synchronous motors at a sacrifice of motor efficiency and power factor correction. Dual drives, that is two pinions driving one gear mounted on the mill, become economical for ball mills drawing more than 3500 to 4000 horsepower (2600 to 3000 kilowatts). Further developments of the low frequency, low speed synchronous motors with the rotor mounted on the mill shell or an extension of one of the mill trunnions could improve the cost picture for these "gearless drives", making them practical for large ball mills. The percent of critical speed, which is the speed at which the centrifugal force is sufficiently large to cause a small particle to ad- here to the shell liners for the full revolution of the mill is given in mill specifications. Critical speed is determined from the following: Where D is mill diameter inside liners (specified in meters). Cs is critical speed in RPM. When D is specified in feet

Jan 1, 1998

By G. Sanders

Discussion by G. Sanders The studies reported on in this paper were initiated to draw attention to the severe contamination problem in the Section 30 drilling program at Chimney Creek. The lithologic-subset sampling study reached a different conclusion from that presented in your paper, and I wish to comment on your subsequent analysis of the data and your conclusions. Request for more complete data In the section on subsampling, you mention that the subordinate lithologies were separated and sampled, yet only the dominantlithology gold value is plotted in Fig. 4. In a contamination study, the reader is interested in the assay values for the individual subsets. Please include a table of the subsample assay data in your reply. Also, please indicate which analytical methods were used to arrive at the gold values in the subsampling study. Turning barren rock into low-grade ore Figure 5 is very revealing and typical of all of the cross sections in Section 30. Note the long strings of low-grade mineralization spread out for hundreds of feet below the ore zones. There were some very high gold values found in certain contaminated fractions during the subset sampling. The conclusion, here, was that the distinctive, strongly-mineralized dolomite layer was probably loose and crumbly and continued to disintegrate during drilling. This caused salting of the unmineralized rock samples below. Missing the high-grade part of the ore body In your statistical analysis, you directly compare the reverse circulation assays to the diamond drill assays in Section 30. Two points argue against a direct comparison and suggest the differences are greater than the 3 % that you report. First, any core loss in a gold zone most likely means that the true gold values are greater. The drillers lost significant amounts of the clay-rich, Section-30 gold mineralization. Also, the initiated salt-mud system, an attempt to improve the core recovery, met with little success. Second, the practice of not sampling core geologically, but instead sampling on even 5-foot intervals, adds a deliberate dilution to the core assay values by including a portion of nonmineralized rock in the first and last samples of each high- grade intercept. The result is often a pair of low-grade assay values on either side of a high-grade gold zone. In reality, a high-grade gold zone has a very sharp assay wall that is often bounded by barren rock. This sampling method may make the diamond-drill core assays more like the reverse circulation values and may help explain the statistical similarities you found. However, it does not represent the true gold values in the high-grade parts of the deposit. You cannot deny that, by careful geological sampling of the drill core, higher and sharper assay values will be obtained. The low core recovery and the diamond-drill-core sampling method used act together to lower the diamond-core assay values. The 3% difference you found between the reverse-circulation and diamond-core assay values could be much larger when you consider what the true diamond-drill core values would be with optimum core recovery and a geologic sampling method for the core. Should statistics have been applied here? The statement "... that reverse circulation holes have overestimated the values of some ore zones and underestimated the values of others" (p. 345) is not correct. The subsampling confirmed what the cross sections hinted at in Section 30. Namely, beneath the high-grade zones, the reverse circulation holes created, by contamination, large intercepts of low-grade ore in regions of barren rock. Because the low-grade material was not there to begin with, this is not a process of overestimating low-grade mineralization. The next statement that "the average result is similar to that of the diamond drill holes" may apply to the data set numerically, but it is not true when viewed spatially on cross sections. Adjacent reverse circulation and diamond drill holes are almost impossible to correlate, high-grade zone values vary widely and many low-grade intercepts make no geologic sense. The subset sampling and cross sections presented in the first part of the paper show that the reverse circulation portion of the data set has some serious problems, as highlighted above, and should not have been dealt with statistically at all. Conclusion Each ore body is different, and each drilling method presents unique sampling problems. In this case, the diamond drill is the

Jan 1, 1994

By J. Lovering, H. Wilhelm, P. Siewert

Introduction The Afton property is located 290 km (180 miles) by air east-northeast from Vancouver and 14 km (8.7 miles) west of Kamloops, a city of 60,000 people, in south central British Columbia, Canada. The mine is adjacent to the Trans-Canada Highway at an elevation of 670 m (2198 ft) above sea level. The ore body is a porphyry copper deposit that has undergone supergene alteration. The major economic minerals in the supergene zone are native copper and chalcocite with chalcopyrite and bornite in the hypergene areas. The grade is 1% with an overall copper distribution - 70% native, 25% chalcocite, and 5% chalcopyrite with bornite and covellite. The ore also contains important but variable amounts of gold and silver. The mill was designed to treat 6350 t/d (7000 stpd). Semiautogenous grinding was selected to minimize capital cost and because of the expected high clay content of the ore, which would have caused problems in a conventional crushing and screening plant. Test work indicated that a recovery of 87% was possible in a circuit incorporating both flotation and gravity separation. Flowsheet Run-of-mine ore is crushed in a 1.06 x 1.65-m (3.5 x 5.4-ft) Allis Chalmers gyratory crusher set at 228.6 mm (9 in.), closed side setting. The surge pocket, below the crusher, is emptied by a Hydrastroke feeder onto number one conveyor, which discharges onto a 180,000-t (198,416-st) coarse ore stockpile. Six Hydrastroke feeders on two conveyors withdraw the crushed material from the bottom of the pile. These two conveyors, in turn, discharge onto the belt feeding the semiautogenous mill. The live storage in the stockpile is approximately 22,000 t (24,250 st), sufficient for three days' mill feed. Primary grinding is accomplished in an 8.5-m (28-ft) diam by 3.7-m (12-ft) long Koppers (Hardinge Cascade) mill (Fig. 1) containing a 10% ball charge and driven by a 4000-kW dc variable speed motor. The mill dis¬charge is pumped by a 10 x 12 G.I.W. pump to a 1.22 x 4.88-m (4 x 16-ft) stationary screen sloped at 20°. Screen oversize returns to the semiautogenous mill (SAM), and the undersize flows by gravity to the ball mill discharge pump box. Secondary grinding is performed in a 5-m (16.4-ft) diam by 8.84-m (29-ft) Koppers overflow ball mill driven by a 3430-kW synchronous motor through an air clutch. The mill is in closed circuit with a Krebs Cyclopac containing 10 635-mm (25-in.) cyclones and the cyclone overflow, at 35% solids and 65% to 70% -200 mesh, is flotation feed. In order to limit the buildup of native copper, circulating in the secondary grinding circuit, a portion of the underflow from the cyclones is processed in a circuit containing screens, cyclones, and shaking tables to produce a finished metallic copper concentrate. Primary mill variable speed drive The overall waste to ore ratio at Afton was 4.5:1. The mining was to be done with only three shovels, which meant that it was highly unlikely that more than one of them would be in ore at any one time. The resulting inability to blend the mill feed made it impossible to prevent wide swings in the grade and grindability. The variable speed do drive motor installed on the semiautogenous mill was selected because of the extreme variability of the Afton ore body. This variability has persisted throughout the lifetime of the mine. There are times, however, when due to ore conditions, the mill is operated at full speed (78% of critical) for extended periods of several shifts duration. There are other times when the mill speed may be changed several times in a 12-hour shift due to changing ore conditions. When ore is processed that contains a fairly large proportion of fine native copper, the primary mill speed and, consequently, the tonnage may be reduced to improve the secondary grind and to maintain an acceptable grind and recovery. High clay ores require less mill speed and more dilute grinding densities. In the latter case, the slower primary mill speed also helps to minimize damage to the mill liners. Approximately 57% of the time the mill operates between 90% and 100% of full speed or between 71% and 78% of critical. The variable speed is also used for inching during mill relines.

Jan 1, 1987

By R. J. Pindred

Discussion by R.J. Pindred In his paper, Davis presents an overview of risk. He also introduces the Capital Asset Processing Model (CAPM) as a foundation for selecting the appropriate discount rate for a mining project. While applying portfolio theory is more defensible than the ad hoc adjustment of discount rates, the CAPM is not a panacea. CAPM shortcomings [The CAPM, as Davis stated, is expressed in the equation: ri=rf+pi4) where ri is the project discount rate rf is the risk free interest rate (3i is the project beta, and 0 is the market risk premium (rm - rf)] Application of the CAPM is more difficult than Davis indicates. Valuation is prospective, while the CAPM parameters are historical. Beta is determined from a regression analysis of historical data, while the beta needed for valuation is the expected beta. Betas are known to be unstable and the regressions that generate them often have low explanatory power. The difficulty of estimating a "project" beta must also be considered. Thus, the beta that is used in the CAPM will be based on the analyst's judgment. Like Cavender's discount rate, this judgment can lead to different project NPVs. Subjectivity in valuation cannot be avoided by a mechanical application of the CAPM. The risk-free rate, which Davis identifies as a short-term real rate of 4%, is also subject to scrutiny. A mining project is not a short-term investment and no single risk-free rate is appropriate for all of the cash flows. The hypothetical mine discussed in Cavender's paper is a six-year project. One might argue for the application of a risk-free rate from the Treasury yield curve at the duration of the project (in a bond-duration sense). This, too, is inappropriate. The risk-free rate should be matched to the timing of the cash flow. These rates can be determined by calculating the implied forward rates from the yield curve using a procedure known as "bootstrapping." It is likely that each of the project's cash flows would be discounted at a different rate. Commodity prices Davis criticizes the "ad hoc adjustment to the discount rate." Yet, in his discussion of the value of stochastic simulation, he suggests that the gold price be modeled as a "random walk, with or without a trend." This is essentially an arbitrary modeling of price risk. Consider that a liquid market in gold futures exists. The futures' price curve, which is closely related to the market's estimate of future spot gold prices, should be used to provide inputs to the model. This is especially true of a relatively short six-year project. Alternatively, as Davis correctly points out, a risk-averse investor can sell the commodity short to hedge price risk. Is it any more correct, in the portfolio sense, to account for price risk at all ?? References Cavender, B., 1992, "Determination of the optimum lifetime of a mining project using discounted cash flow and option pricing techniques," Mining Engineering, Vol. 44, No. 10, pp.1262-1268 Fabozzi, F.J., 1993, Bond Markets, Analysis and Strategies, Second Edition, Prentice Hall, Inc. Higgins, R.C., 1992, Analysis for Financial Management, Third Edition, Richard D. Irwin, Inc. Solnik, B., 1991, International Investments, Second Edition, Addison Wesley Reply by G.A. Davis Pindred discusses two issues related to my paper, the shortcomings of the Capital Asset Pricing Model (CAPM) and which commodity price values to use in the valuation exercise. Even though these topics are not directly related to the use or misuse of Monte Carlo simulation, they are important points to take into consideration in valuation exercises. Since I do not appear to have addressed these issues satisfactorily in my original paper, I will comment on each here. Pindred agrees with me that applying portfolio theory, and specifically the CAPM, to the selection of project discount rates is more defensible than ad hoc methods. But he then points out that the application of the CAPM to project valuation is more difficult that I indicate. It is true that the CAPM is a difficult tool for project valuation in general,. But the application of the CAPM to mining projects is one of the easiest I can think of. The biggest problem with using the CAPM for project valuation is coming up with an expected project beta. I suggest a project beta for gold projects of 0.45. The "true" value might be 0.35, 0.55 or whatever. Pindred correctly notes that the selection of the appropriate project beta is based

Jan 1, 1996

Conveyor belting-Dunlop Belting Division has published a manual on its Starflex plied conveyor belting. The design section of the manual contains advice on the calculation of tensile strength and horsepower needs while the section on belt selection offers helpful recommendations. Circle 200 on reader service card Hydrostroke feeders-A pamphlet from Kone Corp. highlights the uses and operating principles of its hydrostroke feeders. Circle 201 on reader service card Electric cylinders-A 24-page catalog from Raco International Inc. describes applications for its electric linear actuators in addition to the electronic options for computer - controlled operation. Circle 202 on reader service card High torque drives-T. B. Wood's Sons Co. offers a 56-page booklet providing features and specifications on its high torque drives. Information includes a step-by-step drive selection proce¬dure. (HTD) Circle 203 on reader service card Sludge depth meter-The model 600 sludge-depth meter that locates the sludge bed in clarifiers and settling tanks is described in a four-page bro¬chure from Markland Specialty Engineering Ltd. (600-84) Circle 204 on reader service card Cavity pumps-An eight-page bulletin is available from Robbins & Myers Inc. It features the application of Moyno progressing cavity pumps in handling composite slurry fuels. (400) Circle 205 on reader service card Roller chain-A roller chain catalog shows heavy duty drive chains and other specialty conveyor chains. It is available from Peer Chain Co. (PC200) Circle 206 on reader service card Belt filter-Phoenix Process Equipment Co. has available a pamphlet detailing its belt filter press. The unit is designed to dewater refuse and clean coal, yielding easily handled dry filter cakes. Circle 207 on reader service card Capabilities - Literature from International Mineral Services Ltd. highlights its services and capabilities to the mining industry. Circle 208 on reader service card Motor analysis - How to select the proper electric motor by comparing life cycle costs, power costs, rate of return, and other factors, is described in a brochure from Westinghouse Electric Corp. (SA-11376) Circle 209 on reader service card Cavity pumps - A product application data sheet is available from Robbins & Myers Inc. It details the use of Moyno positive-displacement, progressing cavity pumps in handling ground limestone slurry. (PC-21) Circle 210 on reader service card Dust collectors - "Dust Collector Selection Guide," from American Air Filter, describes dry mechanical collectors, wet collectors, fabric collectors, and electrostatic precipitators. (CAD-1-901G) Circle 211 on reader service card Wet scrubber - A 12-page bulletin from The Ceilcote Co. provides a comprehensive description of its ionizing, wet-scrubber system. (12-19) Circle 212 on reader service card Metric o-rings-Simrit Corp. has published a 16-page brochure detailing its full line of standard metric o-rings. Information includes graphics and dimensional charts, and specific data on materials and application ranges. Circle 213 on reader service card Hearing protection - A 16-page catalog from Cabot Corp., EAR Division, provides information on its hearing protection devices and noise control products. Circle 214 on reader service card Toxic gas detection - Sensidyne Inc. is offering a guide for toxic gas monitoring. A description of the electrochemical sensors, as well as ranges, complete specifications, and interference charts are included. Circle 215 on reader service card Hydraulic bolting systems - Ingersoll-Rand Co. is offering a brochure on its line of hydraulic bolting systems. These systems, hydraulic wrench and power console, are designed for heavy duty bolting applications. Circle 216 on reader service card Temperature monitoring - A brochure describing the Ramsey Engineering Co.'s micromonitor temperature monitoring system is available. Three types of switches are available for monitoring bearing temperatures. (80.300) Circle 217 on reader service card Product catalog - Shadbolt & Boyd Co. has published a product catalog. Among items described are hoist slings and cranes; compressors; hydraulics; wire rope, chains, and fittings; and materials handling and shop equipment. Circle 218 on reader service card Cylinder controls - A 12-page booklet presenting Hanna Corp.'s line of electrical controls for cylinders is available. It features proximity and limit switches for hydraulic and pneumatic cylinders, and standard and 3-amp reed switches for pneumatic cylinders only. (550) Circle 219 on reader service card Slurry pump - Pettibone Corp. has published a 24-page booklet covering its heavy duty pumps made with 'diamond alloy' materials for handling slurries of abrasive materials. Circle 220 on reader service card

Jan 9, 1985

By T. J. Jr. DeMull, F. G. Miller, J. P. Matoney

For some time a need had existed in the minerals processing field for a relatively efficient separator that would treat high tonnages of particles in the intermediate size range, i.e., those particles too large for froth flotation and too small for conventional gravity-type separa¬tors. Among those devices developed to meet this need are the centrifu¬gal specific gravity separators. These devices employ centrifugal acceleration to assist gravitational acceleration in separating light¬density minerals from heavy-density minerals. In the category of centrifugal specific gravity separators are the heavy-media centrifugal separator and the water-only cyclone. The two major centrifugal heavy-media separators, i.e., the heavy-media cyclone and the DynaWhirlpool, as well as the water-only cyclone, are discussed in terms of: design features, operating variables, operat¬ing data, and flowsheet design criteria. Examples of plant applications are given in the field of coal processing as well as the processing of other minerals such as iron ores, potash, and tin. Finally, the subject of the staging of centrifugal separators and their use in combination with other separators is discussed. PRINCIPLES For coarse sizes of minerals, efficient specific gravity separations have been possible for many years with open-bath vessels using the natural settling velocity or buoyancy of the particles. These bath ves¬sels process ore by utilizing micron-size solid particles suspended in the slurry fed to the separator. The inclusion of these particles in the slurry increases the effective density of the separating fluid to allow particle separations to be made at densities greater than that of water. However, if vessel size is to remain within economical limits, the particles processed in the bath vessel must have high settling rates in a IG gravitational field. Because of this requirement, heavy¬medium bath vessels are usually restricted to processing +V4-in. sizes. To extend efficient specific gravity separation to smaller sizes, the gravitational acceleration of particles is replaced by centrifugal acceleration. The settling of a small particle in a fluid in a centrifugal force field is similar to that found in a static bath except that the acceleration due to gravity, g, is replaced by a centrifugal acceleration where v, is the tangential velocity at radius r: V=kdm(P-P,), V'. (J) µ In more practical terms where the particles settle in a suspension of finer particles comprising the heavy media and with an effective suspension density p" V = kdm' (P P ) . v'. (2) U r To date the most effective use of this principle has been obtained with devices that rotate a liquid or suspension within a stationary enclosure in order to create centrifugal force. Cyclones are the most common devices used for this purpose, because they generate centrifu¬gal forces far greater than the force of gravity and therefore not only have high capacities but can treat finer sizes than bath-type vessels can. The two main types of cyclones used by industry are the heavy-media cyclone and the water-only cyclone. Also quite widely used is the DynaWhirlpool, which, though based on the same princi¬ple, differs in design from the conventional cyclone. HEAVY-MEDIA CENTRIFUGAL SEPARATORS Like the bath vessels, the heavy-media centrifugal separators em¬ploy media composed of micron-size particles suspended in water. However, the centrifugal force generated in these separators accentu¬ates the difference in settling rate between particles of different density and thus makes possible separations of finer size particles than can be treated in bath vessels. The two most common heavy-media centri¬fugal separators are the heavy-media cyclone and the DynaWhirlpool. Heavy-Media Cyclone Although cyclones were originally developed for use as classifiers or thickeners, it was later found that they could also effectively serve as heavy-media separators.63. 64, Design Features Fig. I I is a schematic of a typical cyclone developed to serve for any one of the following purposes: as a classifier, thickener, or specific gravity separator. The cyclone consists of a cylindrical section joined to a conical section, usually having an included angle of between 14° and 25°. Feed enters the cyclone tangentially through an orifice attached to the cylindrical section. The overflow orifice is located in the base plate of the cylindrical section. The vortex finder, a tube attached to the overflow orifice, extends into the cyclone from the base plate of the cylindrical section. The underflow orifice is located at the apex of the conical section. As some medium together with mineral particles is fed through the feed orifice, a vortex with a hollow air core extending from the overflow to the underflow orifice forms in the cyclone while hollow spray discharges form at each of these orifices. Under the influence of the centrifugal force, high specific-gravity particles move through the medium to the wall of the cyclone and descend in a spiral flow pattern to the underflow orifice. Those particles in the feed stream, lower in specific gravity than the feed medium, follow the major portion of the flow to the center of the core where they are caught in the high-velocity upward central current and are carried out through the overflow orifice. Fig. 12 shows a family of curves that illustrates how materials of varying specific gravity are recovered by a cyclone. Since the specific gravity of the medium is 1.40, the particles of 1.40 sp gr actually act as part of the medium and, regardless of the particle size, split between the underflow and overflow of the cyclone in proportion to the volume split of the medium. Particles higher in specific gravity than 1.40 are recovered in the underflow of the cyclone at increasing rates as the difference in specific gravity increases and the particle size increases. Particles lower than 1.40 in specific gravity are dis¬charged through the overflow orifice at increasing rates as the specific¬gravity difference increases and the particle size increases. However, all the curves originate at the fluid-flow ratio point for the finest particles of any gravity. The fluid-flow ratio is defined as the ratio of the rate of fluid flowing from the underflow to the rate of fluid

Jan 1, 1985