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By Thomas H. Koenning, William E. Bruce

Determination of the state of a mine fire is an important function during firefighting and during rescue and recovery operations. It is imperative that an accurate, up-to-date assessment of fire conditions be available for persons directing such activities in order to avoid unacceptable risk to mine rescue teas and/or fire fighters. Many indicators which consist of ratios of the concentrations of fire gases have been developed over the years for determining the state of a mine fire. Results from numerous gas sample analyses are necessary for determining the trend of the fire. Some indicators are preferred because the influence of sample dilution, a frequent problem, is said to be eliminated. Definitions and examples of indicator6 found to be useful in diagnosing recent western U. S. coal mine fires are presented. Mine gas explosibility is an important subject also requiring rapid assessment. Example explosibility calculations are presented applying one widely used method. Establishing limits for withdrawal of personnel are discussed.

Jan 1, 1987

By George Aul, Richard Ray

Short-hole horizontal degasification systems have been utilized in Island Creek's Virginia Division mines since 1986. These systems remove methane from the coal seam in advance of longwall mining. Without an in-seam methane drainage program, methane liberation rates from the longwall faces may exceed 0.47 cu m/s (1,000 cfrn), requiring 47 cu m/s (100,000 cfm) of air for adequate dilution capability. Optimization of the degas systems has enabled methane liberation levels and the associated ventilation requirements on the longwall faces to be reduced by 50 to 80 percent. Since the inception of the first degas system at the Virginia Pocahontas No. 6 Mine, Island Creek has developed an overall methane drainage program designed to maximize gas production and minimize ventilation requirements for methane dilution. Gas production from the degas systems is affected by four main design criteria: (1) hole location, (2) hole length and spacing, (3) collection system capacity, and (4) time available for drainage. Drilling equipment and materials for the gas collection system are selected to meet the necessary criteria while minimizing manpower requirements. Benefits gained from reducing methane liberation rates from the longwall faces are fourfold: (1) improved safety due to lower methane concentrations in the face area, (2) fewer production delays caused by excessive methane concentrations, (3) lower fan operating costs resulting from reduced air requirements, and (4) enhanced worker comfort resulting from lower air velocities and less dust generation.

Jan 1, 1991

By Nathaniel Arbiter

In attempting to assess the contributions of Antoine Gaudin to the profession which he illuminated for almost a half-century, we will need to examine a very broad canvas. He was active in three significant areas: as a teacher; in research, both through his own endeavors and those of his students; and as a consultant to the mining industry and to the Government. Beyond these specific contributions, there was the major personal influence he had on scores of students, friends, and colleagues in the profession. If a single measure of the man's contribution is to be found, it will be the extent and direction in which he influenced the practice of his profession. To find this we should examine the state of miner- al processing when he began his activities in the 1920's, the progress which has taken place over the past half-century, and the part his activities played in this advance. However, there is a provocative and broader examination which I have chosen to make instead. This is to examine the past century in mineral processing through the lives of three men who, by common agreement, stand on the heights, both in teaching and in their contributions to industry: Robert H. Richards, Arthur F. Taggart, and the man whose memory we honor this evening, Antoine M. Gaudin. Their careers, by coincidence, cover just a century. This is also AIME's first century, and the first century of mineral industry education in the United States. In fact, mineral processing itself on any significant scale in the country encompasses only a century and a quarter. BEGINNINGS OF PRACTICE IN THE U.S. Before the 1840qs, U.S. mining, except for coal and iron ore, was on a small scale and was confined to the eastern states, with processing practically nonexistent. The 1840's saw the beginning of Michigan native copper production, including ore dressing; the 18501s, iron ore production in Minnesota; and the 18601s, lead production and processing in Missouri. But the most explosive event was the California gold rush after 1848, which not only brought tens-of-thousands of fortune hunters to the West, but also catalyzed broad mining activities and processing within the following decades. This was not only in Calif-

Jan 1, 1976

By E. J. Wasp

An upsurge of interest in slurry pipelines to move coal from Wyoming, Montana, and Colorado to markets in the southwest, west, midwest and northwest has recently developed. Specific pipelines have been proposed to serve these markets; the routes of these lines are depicted in Exhibit I. It is, therefore, timely to discuss some technical aspects of the terminal facilities for these western coal pipelines. The discussion will be based on experience with the Ohio and Black Mesa pipelines as well as detailed planning and testing for a proposed 1,000-mile system from Wyoming to the middle south area. The economics of the pipeline terminal processes are heavily depend¬ent on site, type of process used, and other specifics of the situation. However, the following approximate breakdown of costs for a 1,000 mile, large capacity pipeline will be useful for orientation purposes. [Water supply 4 Coal cleaning 5 Feed preparation 10 Slurry transportation 72 Dewatering terminal 9 100%]

Jan 1, 1976

By D. A. Elgillani

Oxidation potentials have been measured in the presence of various concentrations of cyanide, ferrocyanide, and ferricyanide and ethyl xanthate at various values of pH and related to flotation response. Eh-pH diagrams are constructed and show that the formation of surface ferric ferrocyanide is probably responsible for depression when cyanide is added. The influence of cyanide in the depression of pyrite with xanthates as collector has been the subject of many investigations (1,2,3,4,5,6), and from these studies several theories on the mechanism of depression have evolved. Wark and Cox (7) and Gaudin(8) have suggested that the depressing effect is due to a competition of cyanide ion with xanthate ion for the surface. Cook and his colleagues (9,10,11) have explained this phenomenon in terms of competition between hydrocyanic acid and xanthic acid. Sutherland (12) has shown that both of these theories describe accurately the relation between pH value and cyanide addition at constant collector addition but fail to describe the relationship, between pH value and the amount of collector required to cause flotation. Taggart(12) suggested that depression in these systems is due to the formation of a reaction product between ferric ion at the pyrite surface and ferrocyanide ion derived from solution. Majumdar(4,6) has attempted to prove this hypothesis by measuring the contact angles of pyrite in the presence of 25 mg/liter ethyl xanthate and different concentrations of potassium ferrocyanide and ferricyanide. In all cases the contact angles were quite high up to pH 10; these results indicate that pyrite should not be depressed by either potassium ferrocyanide or ferricyanide. In view of these facts, Majumdar has assumed that the compound Fe(CN)2 forms at the surface.

Jan 1, 1968

By R. Mensah-Biney

The kinetics of loading of gold from bromide solution onto an anion exchange resin, PAZ-4, were investigated in a batch stirred reactor. The initial loading rate was described mathematically by a first-order reaction rate equation. The effect of agitation speed, temperature, resin particle size, resin concentration, initial concentration of gold and bromine in solution on the rate was investigated. The model described fairly accurately the loading process in terms of the variables investigated. The rate constant for the initial loading ranged from 0.5 to 1.3 x10-3 sec-1 and the half-time was 9.3 to 56 minutes. The initial rate constant increased with increasing temperature with an activation energy of 15.7 kJ/mole (3.8 kcal/mole). The rate constant was directly proportional to the concentration of the resin in solution. The subsequent reduction reaction was also modeled by the first-order rate equation and the apparent rate constant was estimated to be 3.0 x 10-6 sec-1 with a half-time of about 44 hours.

Jan 1, 1994

By R. C. Bunge

Many minerals float best with fatty acids in the intermediate pH range. For example, fluorite, a semi soluble salt mineral, exhibits a maximum contact angle and a maximum flotation recovery in the pH range of 6-8 after conditioning in a palmitate solution (fig. 1). The maximum in flotation coincides roughly, as many researchers have observed, with the boundary of bulk precipitation of the fatty acid. According to fig. Z this boundary is in the above case of 5xl0-7 M palmitate approximately at pH 6.3. For assessing the phenomena occurring at the bulk precipitation boundary of fatty acids, the surface tensions of carboxylate solutions have been measured under various conditions. Fig. 3 shows the surface tension development of palmitate and myristate solutions of increasing concentration at pH 12. Also shown are the bulk precipitation edges (BPE) of the respective solid species that is stable under these conditions, that is the potassium acid soap, HKAZ. Both curves display a similar behavior. As the concentration of carboxylat,e increases, the surface tension decreases, until a minimum is reached at which the curves level off. The decrease of surface tension with increasing carboxylate concentration is generally taken to indicate the surface excess of the carboxylate ion. The point where the curve suddenly levels off has not yet been explained. While it is well known that the onset of micellization (CMC) may cause such an inflexion of the surface tension curve, this cannot be the case here because the position of this point depends, as shown in fig. 4 for palmitate, on pH (the CMC is generally independent of pH). We believe that this point indicates the precipitation of a colloidal acid soap (colloidal precipitation edge, CPE) before the bulk precipitation of this species sets in (BPE).

Jan 1, 1994

By Richard L. Bullock

OPEN STOPING An open stope is an underground cavity from which the initial ore has been mined. Caving of the opening is prevented (at least temporarily) by support from the unmined ore or waste left in the stope, in the form of pillars, and the stope walls (also called ribs or abutments). In addition to this primary support system of open stoping, some secondary support may also be required using rockbolts, reinforcing rods, split pipes, or shotcrete to stabilize the rock surface immediately adjacent to the opening. The secondary reinforcement procedure does not preclude the method classified as open stoping. There are many forms of open-stope mining used to extract the initial material from a mine. Having once established that the mineral and waste rock are competent enough to use an open-stoping method, and assuming that the reserve is not classified as gassy, the Form which the method will take is primarily determined by the dip and thickness of the reserve. How these two factors affect the selection of the open-stope mining is discussed in a later chapter. At this point it will suffice to say that the classification of the open-stopes mining system which follows is based on whether dry' broken materials flows by gravity or whether it must be moved by nongravity methods where energy must be supplied to move the material. Room-and Pillar-Mining Room-and-pillar mining is an open-stoping method where mining progresses in a nearly horizontal or low angle direction by opening multiple stopes or rooms, leaving solid material to act as pillars to support the vertical load. Since the direction of excavation (angle of dip) is below that which would cause the dry material to flow by gravity to a drawpoint or gathering point, the material must be loaded in the room where it was extracted and transported to a point where it will flow, either by gravity or mechanical means, to a central gathering point to be taken out of the mine. This is an important aspect of room-and-pillar mining which differentiates the system from other open-stope mining methods which rely heavily upon gravity to transport ore from where it was broken to a lower elevation, usually through a drawpoint. There are many variations of the method which go by a number of names in local districts: breast stoping, breast-and-bench stoping, board-and-pillar, stall-and-pillar, and panel-and-pillar are all basically open-stope room-and-pillar mining. In some instances detailed stope planning is almost nonexistent; i.e., the operator simply follows the visual pay values, leaves pillars only where necessary, and tries to locate them in the zones of lower value. This method of mining is as old as the beginning of underground mining itself, dating back thousands of years. Early in the history of mining in this country, the term "gophering" was used to describe this method (Peek 1941). The term is appropriate, for it brings to mind the exact results of this type of system-a random and irregular room-and-pillar mine. In other instances where the mineral values are consistent both in physical dimensions and quality, the mine layout can be planned to the last detail, resulting in a uniform room-and-pillar mine. Coal, trona, gilsonite, potash, oil shale, salt, limestone, and sandstone mines can usually follow such a system. Today, most metal mines using a room-and-pillar operation try to mine as regular a pattern as possible but deviation in height, width, thickness, dip, and grade of the ore results in comparable deviation in the mine plan. Variations of the Room-and-pillar System It is necessary to briefly describe some of the many variations of the room-and-pillar system of mining, enabling the reader to fully explore the concepts and become familiar with the terminology used before going on to the details of mine design. Full-Face Slicing: If in the process of opening the rooms the total vertical extent of the mineral values of the particular seam or strata are extracted from the advance of one vertical face, the term used to describe this is full-face slicing. This face is also known as the "breast." There is no mineral of economic value intentionally left either in the floor or the roof (back) to be mined later. To be able to extract the full-face height in one pass, the mining equipment must obviously be designed to reach as high as the back. In an Appalachian coal mine, this may be all of 660.4 mm (26 in.); for a future oil shale property it might be 15.24 m (50 ft). Normally, however, in a majority of mines where the mining face gets over 6.09 to 6.7 m (20 to 22 ft), the tendency is to divide the face into more than one pass. Over this height, it becomes difficult to properly see and remove loose rock from the back with a hand "mining bar." Where the process of taking down loose rock has become mechanized, higher full-face mining can be safely practiced. Most eastern and midwestern coal seams and western uranium, trona, and potash seams in the United States are easily reached in a single face; many limestone, lead, and zinc mines must resort, at least in part, to "multiple-slicing" to remove all the minerals of value. Multiple Slicing (also known as multiple-pass mining): In many cases it is not practical to carry the full vertical height of the mining horizon as a full face. The face is divided into parts known as the breast, bench, and/or brow. Ideally, if the operator knows the vertical extent of the mineralized zone, he will drill and blast the first pass at the top of the zone, thereby

Jan 1, 1982

By Stuart Ratcliffe, Denis Kidby

The development of Vitrokele™ Technology over the last 10 years is reviewed with reference to mine site piloting and com­mercial applications. The first mine site piloting was conducted in Australia in 1989 followed by a further six continuous site trials in Australia, Canada and the USA. These major pilot plant studies are complemented by over 50 laboratory scale batch and pilot testwork programmes. The first commercial plant using Vitrokele™ Technology was the Connemara project commissioned in Zimbabwe in 1995. Subsequent to this, three commercial scale projects have been designed using Vitrokele™ to recover precious metals, base met­als and cyanide. The various projects are discussed with relation to the differ­ing flowsheets applied to each particular case.

Jan 1, 1998

By Slavoljub D. Maksimovic, Steve J. Page

The Bureau of Mines measured respirable dust emission rates from overburden drilling at a surface coal mine and the transport of this dust to downwind locations. The emission rates were estimated by using multiple gravimetric dust samplers mounted on a cross-shaped tripod located immediately downwind of the drill. The transport of this dust was measured at more distant downwind locations by a combination of respirable gravimetric sampling and integrated bag sampling of sulfur hexafluoride (SF6), which was used as a tracer gas. Two modes of sampling were used: (1) At five sampling locations arranged in a semicircle around the drill. The radius of this arc ranged from 15.85 m (52 ft) to 71.32 m (234 ft). (2) In-line sampling at downwind locations with samplers located at intervals of either 10.67 m (35 ft) or 15.24 m (50 ft) for a total distance of 76.20 m (250 ft). Samples were taken 1 .22 m to 1.52 m (4 to 5 ft) above ground level. Test results for the semicircular test arrangements showed that at any downwind location, the maximum contribution of respirable dust due specifically to the drilling operation was 42%; the average for eight tests was 13.6%. In general, there was no correlation between the downwind distance of this maximum contribution and the maximum contribution itself because of changing wind direction. Test results for the in-line arrangements indicated that the transport of the dust from the drilling operation usually decreases rapidly with distance and that, even though the drill dust may produce significant health hazards for the drill operator and drill helper (or other personnel in the immediate vicinity), the exposure of downwind personnel to drill-generated dust may not be a significant problem.

Jan 1, 1987

By S. K. Kawatra, H. J. Haselhuhn

"Fine-grained hematite ore can be concentrated by the process of selective flocculation and dispersion, which relies on proper reagent selection and control of water chemistry. While many previous studies have been performed analyzing the effects of different reagents on this process in a laboratory environment, this paper focuses on investigating the water chemistry within the process in a pilot-scale continuous deslime thickener.The pH, sodium concentration, calcium concentration and magnesium concentration were varied to determine their effects on the iron concentrate grade and recovery, and the phosphorus concentrate grade and rejection in the pilotscale selective deslime thickener. The ideal pH for the iron grade and recovery of the process using a starch selective flocculant was found to be 10.5. Phosphorus rejection, however, was increased at lower pH values. Minimization of sodium concentration was shown to improve iron grade, iron recovery and phosphorus rejection. Calcium acted as a nonselective flocculant showing higher iron recovery, lower iron grade and lower phosphorus rejection with increasing concentration. Conclusions could not be drawn from the experiments that varied magnesium concentration.The zeta potential of the solid-liquid interface of particles in each sample taken was also analyzed to show relationships between zeta potential and process performance. In all cases, a maximization of the magnitude of zeta potential correlated with increased iron grade and recovery. This supports the hypothesis that a higher level of dispersion enhances the selective flocculation and separation process.IntroductionThe United States produced 53.2 million metric tons out of the 3,000 million metric tons of concentrated iron ore produced worldwide in 2012 (Tuck, 2013). With this in mind, there has been a large increase in iron ore processing research (Bolen, 2014; Carlson and Kawatra, 2008, 2013; Halt and Kawatra, 2014; Halt et al., 2014; Haselhuhn et al., 2012a,b, 2013a,b; Liu et al., 2014; Manouchehri, 2014; Sandvik and Larsen, 2014; Semberg et al., 2014). Of the 53.2 million metric tons produced, roughly 8 million metric tons of hematite (Fe2O3) concentrate were produced using a process known as selective flocculation and dispersion (Cliffs Natural Resources, 2011). This process is the only economically viable method for concentrating low grade (< 40% Fe) fine grain (< 25µm liberation size) hematite ores. It is scarcely used due to difficulty in controlling the surface chemistry of the ore and high reagent costs (Department of Energy, 2001; Haselhuhn et al., 2013a). The process is highly dependent upon both reagent selection and precise control of the water chemistry during separation. This study shows the dependence of the process performance of a pilot-scale deslime thickener on the water chemistry in the feed slurry."

Jan 1, 2015

By Edward H. Opitz

Much has been written recently about the Henderson Mine Project in Colorado being developed by American Metal Climax, Inc. To briefly recap the project, AMAX has an ore body, containing about 300 million tons of proven and probable molybdenum ore which is located under Red Mountain on the eastern slope of the Continental Divide. The mill, however, will be located on the western slope in the Williams Fork Valley where water and space for waste disposal are more available. AMAX will carry the ore by remote controlled trains from a gathering level underground at the mine, through a ten (10) mile tunnel under the Continental Divide and thence along a five (5) mile surface railroad to the mill site. One of the more challenging facets of the project is the western portion of the ten (10) mile tunnel being driven by Dravo Corporation through difficult ground conditions, downgrade, toward the Continental Divide. The tunnel is a 16&apos;-6" wide by 15&apos; high horseshoe, which runs in a straight line and drops at a constant three percent (3%) grade from the West Portal to the ore zone. Dravo&apos;s portion is anticipated to be about 6 1/2 miles long and includes permanent pumping stations located every 4,900 linear feet (lf) permanent rectifier stations every 12,5001f, and one (1) 1,560 ventilation shaft. The western portion of the tunnel lies in the Idaho Springs formation, a mixture of gneiss and schist with some granitic intrusions. Toward the Continental Divide a transition into Silver Plume Granite is expected to occur. Preconstruction geologic information indicated the Idaho Springs and Silver Plume formations to be basically competent with some faulted and altered areas. It

Jan 1, 1974

By David Brown, Adrian St. John, John Comins, Mike King

"INTRODUCTION The Crossrail project provides an East-West link across London enabling an additional 200 million journeys each year linking up major business centres and existing transportation interchanges, while reducing congestion at street level and within the current underground system. The project has built major structures in Sprayed Concrete Linings (SCL), including 5 new stations, 6 crossover caverns and numerous passages, adits and ventilation tunnels. Construction has taken approximately 4 years, and has involved the spraying of over 200,000m3 of concrete. At the start of the project there was a lack of available records from previous projects concerning concrete falls, which hindered risk assessments and the ability to make meaningful comparisons around spraying or material performance. This paper redresses this lack of data by providing information on falls experienced on this project and makes recommendations to assist future projects. Construction Methods Table 1 below summarises the main contracts with elements of permanent SCL. The listed primary lining lengths include pilot tunnels, side drift tunnels and enlargements as separate tunnels.A range of construction sequences were used in order to deal with the design requirements, geometry, and the ground and ground water conditions. The following briefly summarises the build up by layers, with typical design dimensions provided for a platform sized tunnel (approximately 9m internal diameter), excluding tolerance allowances. Some linings were sprayed in more than one layer to allow for inclusion of reinforcement layers or to avoid a too aggressive build-up of thickness: •,Initial shotcrete, or sealing, layer (ISL) of 75mm followed by a Primary Lining (PL) of 300mm. •,Regulating layer of 40mm followed by a Waterproof membrane (Sprayed membrane, 3-6mm). •,Secondary lining (SL) of 250mm with a separate sprayed fire protection layer (FPL) of 50mm."

Jan 1, 2016

By J. Hahn

Modern high-performance disc filters such as the BoVac Disc Filter have set new standards in applications where large capacities of solids, liquids and air have to be managed through the filters. These disc filters have an enormous hydraulic capacity. For applications in mineral processing - with a special focus on tailings – which typically have high solids throughputs but low filtrate and air flows, this capacity can be used to expand the filtration area while maintaining the benefits of very good cake discharge even with thin cakes (4-5 mm) and a high filter speed. Research work over the last century has proven that increasing the filter area should be achieved by increasing the disc diameter. Because adding more discs leads to big problems with the cake discharge. However, as the disc diameter increases, the filter segments become longer and heavier. A suitable compromise was therefore found by doing both. The result is the first vacuum disc filter that offers significantly more than 300 m² of filtration area, but has no more than 6 discs.

Feb 1, 2023

By J. DuCarme, A. Smith, M. Reyes

"A magnetic proximity detection system is mounted on a mobile mining machine to prevent underground workers from being pinned or struck by machine motion. The system generates magnetic fields around the machine to determine safe working distances. The minerworn component measures the magnetic field in order to approximate location. Large masses of steel, such as those from mining equipment, can alter the magnetic field distribution. This affects the locational accuracy of the system, thus adversely impacting worker safety. To examine this problem, U.S. National Institute for Occupational Safety and Health researchers developed a method and test system to study the influence of a steel mass on the magnetic field distribution. The results show that a steel plate can strengthen the magnetic field perpendicular to the generator by up to 40 percent. Furthermore, they show that the degree of the influence on the field distribution is a function of distance. The results from this study can be used to further develop and improve the performance and reliability of electromagnetic proximity detection systems used in underground mining applications.IntroductionOn average, 250 miners are injured every year, and 43 miners were reported to be killed by striking or pinning accidents involving continuous mining machines in the United States between 1984 and 2015, according to the U.S. Mine Safety and Health Administration (MSHA, 2015). Continuous mining machines are large mobile machines used to mechanically cut coal in underground mines. These machines weigh approximately 60 t (132,000 lb) each. Although modern continuous mining machines can be remotely controlled, operators still need to be close to the machine primarily because of the low visibility and wide blind spots of the underground working environment.The MSHA promulgated a final regulation in 2015 that requires the use of proximity detection systems on all continuous mining machines except fullface machines. MSHA has approved five such systems to be intrinsically safe for use in underground coal mines: (1) Strata Worldwide’s HazardAvert System, (2) Frederick Mining’s Model 200 HazardAvert Detection System, (3) Nautilus International’s Coal-Buddy System, (4) Matrix Design Group’s M3- 1000/Joy Global SmartZone Gen 1 and (5) Matrix Design Group’s IntelliZone/ Joy Global SmartZone Gen 2."

Jan 6, 2018

By Douglas Ball

An underground storage field is a gas field where man, not nature, has put all or most of the gas. All but one of those in use today are in the pore spaces or "porosity" of sedimentary rocks, the most common being sandstone. The one exception, near Newcastle, England, is a gas storage in a cavern dissolved from rock salt. The tens of millions of barrels of LPG (liguified petroleum gas) now stored in salt caverns mined in shale is liquid storage, not gas storage, and in this paper I recognize the presence of these installations but do not include them in the gas storage category, Since, with the English exception, gas storages are gas fields, they can all be sorted into the categories used to describe gas fields, 95% of them once gas fields anyway, but some of the remaining 5% are the ones having the greatest effects on underground waters.

Jan 1, 1960

By Don S. McCoy, W. Joseoph Schlitt, Donald B. Swanson

Introduction Contract miners have long played an important role in the minerals industry. Historically, these were individuals who worked as underground miners and were paid by the ton or the day. These hardy types still ply their trade in the industry. However, today the term "contract miner" generally refers to an independent contractor who supplies the people and equipment needed to conduct surface mining operations for an owner or lease holder. At one time or another, contract miners have probably handled virtually every mineral commodity from sand, gravel, and stone quarrying to mining uranium and precious metal ores. In today&apos;s market, though, contract mining appears to be concentrated in just a few areas. These include precious metal mines, various types of solid fuel operations, and reclamation of abandoned mines. This article covers market areas, contract mining organizations and services, and their advantages and disadvantages. Gold and silver mines Although exact figures are not available, contract mining represents a significant element in the precious metals segment of US mining. For gold and silver, the growth in contract mining appears to be paralleling the overall growth in mining activity itself. Nevada is the largest gold producer in the US. The latest available data (1985) on open pit gold/ silver mines is summarized in Table 1 (Nevada Division of Mine Inspection, 1986). These figures show that contractors mined nearly half of the ore in 1985. Total tonnage mined, of course, was much higher for both owners and contractors. Assuming an average stripping ratio of 2.5 to 1, the total daily mining rate for the contractors was more than 145 kt (160,000 st). This translates into a contract mining rate of about 36 Mt/a (40 million stpy) based on 250 operating days a year. This approximate 50-50 split between owner and contractor also appears to hold for the future. Metals Economics Group (1987) published a survey of 51 short-range and 54 long-range precious metal projects being considered throughout the US. Contacts with owners, requests for proposals, and past corporate philosophies all suggest that about 55% of the open pit operations scheduled for the near-term will be contract mined. This represents an estimated 26 kt/d (28.5 stpd) of ore. A similar assessment of the long-term projects indicates that about 12 kt/d (13,000 stpd), or just under 50% of the open pit tonnage will be contract mined. There are several reasons why gold and silver represent such an attractive opportunity for contract miners. One is that many properties are owned or leased by individuals or partnerships that are essentially entrepreneurial. Many of these owners or lease holders have minimal mining experience. They must therefore either rely on contract mining organizations, sell an interest in the project to another mining company that will become the operator, or hire an experienced management team. For about the same reason, junior mining companies also represent good prospects for contracting. This is because most junior companies focus on exploration. They would prefer to spend their generally limited funds on drilling programs. Thus, they tend to minimize additional capital requirements by using contract miners or farming out the property to another mining company. Another good reason for using contract miners is the size of the deposit. Many properties are so small that their reserves will not support the capital expenditures

Jan 12, 1987

By J. A. Finch, J. Yiantatos, A. R. Laplante

This paper re-examines the mixing conditions in the collection zone of flotation columns. These have in the past been characterized with a single uniaxial dispersion coefficient, E. The relationship between E and column diameter, Dc, is examined based on residence time distributions (RTDs) in columns of Dc ranging from 5.lcm to 184cm. Results agree with published correlations in chemical engineering, i.e. E a Dc &apos;" , Data explain why pilot columns display significant upgrading between the feed level and the interface, whereas plant columns do not. It also suggests that, for plant columns, the full collection zone (up to the interface) can be considered when calculating the residence time. A mathematical model considering two regions of different mixing was compared to the single E model. Both models adequately fitted measured age distributions inside the column and overall RTDs. Measured and simulated mixing characteristics are compared to those of flotation banks, using a tanks-in-series approach; banks of comparable mixing to columns are identified.

Jan 1, 1988

By J. M. Scharer, Bruce E. Halbert, Pettit. Carol M., Lysane Bolduc

The ratio of neutralization potential (NP) to acid production (AP) ratio is often employed for predicting acidic discharges from waste rock dumps. The available NP was evaluated in laboratory studies of limestone dissolution in simulated acidic mine water. The available neutralization potential strongly depended on the particle size. Using limestone particles of 114" (6.4 mm) or greater, less than 20% of the total neutralization potential was exhausted at the onset of acid conditions. Apparently, the dissolution of limestone became mass transfer limited. Under mass transfer limitations, the rate of pyrite oxidation may exceed the rate of neutralization by buffering minerals. The implications concerning the controls on limestone dissolution were examined using a geochemical predictive model for sulphide mineral oxidation in waste rock. The results have shown that in case of heterogeneous waste rock piles the NP/AP ratio is a reliable indicator for short-tam predictions only. Kinetic data on the depletion rate of the neutralizing minerals and geochemical modelling suggest that waste rock environments with NP/AP ratios as high as 5.0 may turn acidic in the long term.

Jan 1, 2000

By Janusz S. Laskowski

Studies carried out over the last 10 years reveal that precipitating particles which may appear in alkaline solutions of long-chain amines and in acidic solutions of fatty acids exhibit clear isoelectric points around pH 10.6-1 1 for dodecylamine and around pH 2.5-3 for fatty acids. These colloidal species can accumulate at solid/liquid and liquid/gas interfaces and affect flotation. While the precipitate in solutions of fatty acids apparently can chemically interact with the mineral surfaces and is flotation-active if sufficiently long conditioning time is available, the cationic flotation of quartz in alkaline solutions is depressed by the colloidal precipitate. The experimental data available in literature on the adsorption, electrokinetic, contact angle and surface forces measurements in the solutions of the weak electrolyte collectors are analyzed and the role played by the precipitate is discussed. It is shown that these concepts allow for an explanation of the selective flotation of sylvite from halite with long-chain amines in potash ore flotation.

Jan 1, 1999