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By Robert S. Hedin, David M. Hyman

Acid mine drainage (AMD) is a common water pollution problem on active and abandoned coal mine sites in the eastern United States. AMD forms when surface mining brings unweathered pyrite-containing rocks to the surface or when deep mining allows oxygen to contact buried pyritic strata. In the absence of neutralizing compounds, the drainage that results can be extremely acidic and contaminated with dissolved iron, manganese, and sulfate. Drainages with pH < 3.0 and concentrations of sulfate greater than 1,000 mg/L, iron greater than 50 mg/L, and manganese greater than 10 mg/L are common. Where water flows through alkaline materials (such as limestone) before surfacing, the drainage is less acidic and occasionally circumneutral, but it can still contain high concentrations of sulfate and metals. Current water quality standards in the United States require that mine discharges have a pH between 6 and 9, total iron concentration less than 3.0 mg/L, and manganese less than 2.0 mg/L. At thousands of active and inactive mine sites, drainage does not meet these standards and is being treated before discharge by the mining company. At thousands of other sites, which were abandoned prior to the enactment of water pollution laws or were operated by companies that have gone bankrupt, untreated AMD is polluting receiving water systems. The standard mine drainage treatment system involves the addition of a1 kal ine chemicals to the water, which raises the pH and cause metals to precipitate in a settling pond. These systems are expensive, often costing tens or hundreds of thousands of dollars per year for chemicals, operation, maintenance, and disposal of the metal- laden sludge. Because the drainage on many sites will likely be contaminated for decades, there is financial incentive to find alternative water treatment systems. In the last five years, many mining companies and engineering firms have experimented with the construction of wetland treatment systems. Very few of the several hundred wetland systems that have been built are performing well enough to justify total abandonment of chemical treatment. Nonetheless, systems continue to be built because the mining companies have found that wetlands can reduce their AMD treatment costs. It is also their hope that further research and development will eventually result in systems that will completely rep1 ace chemical treatment and offer long term performance at minimal cost. The Bureau of Mines has been actively involved, from a research perspective, with this approach to AMD treatment since its inception. In this paper, the status of constructed wetland technology is discussed with respect to the construction and performance of systems, chemical and biological processes that affect AMD chemistry within constructed wetlands, and the future of this technology as perceived by the Bureau of Mines. ORIGINAL OBSERVATIONS The constructed wetland concept has its roots in observations of natural Sphaqnum peat wetlands that received acid mine drainage and, instead of being adversely affected, appeared to cause an amelioration of the polluted water (Huntsman et al. 1978; Wieder and Lang 1982). These observations instigated the idea that wet1 and systems might be used for the intentional treatment of mine drainage. Because the discharge of AMD into a natural wetland is prohibited in the U.S. by several laws, it has been necessary to construct wetlands which act solely as water treatment systems. Initially, most wetland research and construction efforts mimicked the original observations by utilizing Sphaqnum moss and peat. Despite promising laboratory results (Kleinmann et al. 1983; Tarleton et al. 1984; Burris et al. 1984; Gerber et al. 1985), virtually all field tests of Sphasnum-dominated constructed wetlands failed to provide sufficient water treatment for more than several months. Sphasnum proved quite sensitive to the stresses associated with transplanting, abrupt changes in water chemistry, excessive or insufficient water depth, and

Jan 1, 1989

By Timothy C. Smith, S. Andrew Jacobs, Douglas E. Gonzales, Sandy J. Schexnailder, Carl Urban

At the Schwartzwalder Mine (Cotter Corporation, Golden, Colorado) ion exchange resins have been used in mine water treatment experiments. The experiments, initiated in October 1979 and currently in progress, demonstrate that mine water uranium values can be reduced to levels as low as 10 ppb. Duolite® A-101D, a strongly basic anion exchange resin, was tested in laboratory-scale columns (1 inch diameter) initially and then in pilot-scale columns (4 inch diameter), using both co-current and counter-current operational modes. Observations and conclusions resulting from these studies can be summarized as follows: (1) Duolite® A-101D appears to have successfully extracted uranium at influent levels of approximately 2.0 ppm over multiple loading and elution cycles, consistently producing effluent concentrations as low as 10 ppb uranium. (2) Based upon present test data, counter-current regeneration appears to be the key to successful operation. Upflow loading and downflow elution are the preferred modes for service and regeneration, respectively. (3) Filtration is a necessary pretreatment to remove fine clay and inorganic precipitates that will foul the resin, causing excessive pressure drop and poor loading and elution kinetics. (4) This waste treatment application is unique in that it offers the opportunity to recover some or all of the capital investment in the treatment facility

Jan 1, 1980

By Y. H. Alghamdi, A. D. S. Gillies

"INTRODUCTION Background Diesel Particulate Matter (DPM) has been a critical issue during the last decade in underground mines. This problem is threatening the health of mine workers, as well as the mine’s productivity. It has been confirmed that the diesel engine exhaust is the main source for the diesel particulate matter (DPM), and that it has the highest impact in raising carbon monoxide, carbon dioxide, oxide of nitrogen, and hydrocarbons in the underground environment. Miners who are exposed to this kind of emission are at high risk for developing lung cancer and other lung diseases. The Mine Safety and Health Administration (MSHA) has issued standard rules to regulate diesel emissions in underground mines. The final ruling confirmed in 2008 that the total carbon should not exceed 160 µg/m3. (MISHA, 2008) Monitoring DPM in underground mines has become important to ensure compliance with the MSHA regulations. There are many different types of DPM monitoring systems that are currently being used in underground mines. Two methods are typically used for DPM sampling: the personal sampling method or the direct exhaust sampling method, which takes a direct sample from the engine’s tailpipe. Many mines do not recommend using the direct exhaust sampling method since it involves complex measuring equipment and procedures, as well as it is not regulated in the United States or Canada. Thus, the personal sampling method is the preferred method in many mines when taking DPM samples. Even this method includes many variations, such as the Respirable Combustible Dust (RCD) method and the method devolved by the National Institute for Occupational Health and Safety NIOSH 5040. The NIOSH 5040 method is the preferred method overall because it gives more accurate measurements compared to the other methods. This method does not allow organic carbon sources other than those coming from diesel engine exhausts to interfere with the measurements, which otherwise would lead to overestimating the actual amount of the DPM in the mine environment. An impactor feature in the cassette, which holds the samples in the instrument, characterizes this method. Since the diesel particles are less than 0.9 micron, this impactor does not allow particles more than 0.9 microns to be collected in the cassette filter. The cassette is then analyzed in a lab and the concentrations of organic carbon (OC), elemental carbon (EC), and total carbon (TC) are determined. Lately, the National Institute for Occupational Safety and Health (NIOSH) has developed a new instrument that can use the same concept of NIOSH 5040, but with instant measurements of elemental carbon. It utilizes laser radiation with the cassette so it can give a real time measurement. The real time measurements give instant information about the underground environment that will allow the mining engineer to take immediate action and enact plans instead of waiting for results from the laboratories where NIOSH 5040 method may take several weeks to get the measurements."

Jan 1, 2016

By V. Srivastava, T. Ghosh

"Modeling and Simulation of SAG Mill has been widely used in design and optimization of mill performance in terms of its power draw, processing capacity and product size distribution. However, these models are solved under steady approximation and do not provide any information on mill charge distribution in real time. This paper attempts to characterize mill contents by solving Whiten’s first order contents based model in Matlab/Simulink environment. The parameters in the model such as breakage rate (S), discharge rate (D) and appearance function (A) were estimated utilizing process and design data collected from a gold mine operating in Alaska coupled with nonlinear parameter estimation scheme. This model is then utilized to predict dynamic response of other key operational variables such as mill power, bearing pressure, charge level and product size distribution. The transient response of mill behavior with respect to changes in feed size distribution, tonnage and mill feed water is also presented. This dynamic simulation approach can be used for practicing different control strategy and training purposes. The observed response of mentioned variables was validated utilizing dynamic response data from the plant to encompass operational characteristics of SAG Mill behavior. Keywords: SAG Mill, Dynamic Modeling, Simulation, Nonlinear Parameter Estimation. INTRODUCTION Grinding in mineral industry is one the most important steps in the process of size reduction and plays a crucial role in determining the overall economics of the milling plant. Much of the efforts focused on improving the process efficiency using modeling and simulation have been practiced by researcher and engineers. However, up till now these models were simulated under steady state approximation because of their complex nature and difficulty in programming them in a dynamic framework. Steady state simulations have been successfully utilized in circuit designing, capacity planning and pre-flowsheet optimization. The maturity in steady state models is because of its applicability and regular use in steady state simulation packages such as JKSimMet, Limn, ModSim, etc. With the advent of equipment such as SAG and AG mill has been a recent shift in the simulating these models using dynamic approach. Because of the economic benefits of these devices and ability to act both as a primary grinding device and as secondary or tertiary crushing device they have gained a large popularity and are now universal to any milling plant. However, these device poses a real challenge to engineers and researchers in terms of operation and control. The major reason attributed to problems faced is because the ore constitutes the major portion of the charge in these devices when compared to ball and rod mill. This rock charge is subjected to disturbance such as changes in feed size distribution and ore hardness which results in unstable operation and frequent mill stoppages. One way to prevent this is to measure and control these external variables, however, lack of available instrumentation and unavailability of sensors to measure these variables has foiled the understanding of complete process occurring inside the mill thereby, inhibiting development of proper control system."

Jan 1, 2017

By H. L. Ehrlich

Marine manganese nodules harbor a mixed population of manganese-oxidizing and -reducing bacteria, as well as bacteria that do not act on manganese. They also host phagotrophic protozoa. Out of several laboratory studies, one line of investigation indicates that some kinds of bacteria found on nodules can catalyze the oxidation of Mn2+ bound to Mn(IV) oxide at the surface of nodules at in situ temperature and pressure. In the absence of appropriate bacteria, the rate of oxidation of the surface-bound Mn2+ is significantly slower. How Fe(III) is incorporated is still unclear. Base metals such as Cu, Ni and Co, which are usually found in marine nodules in various quantities, may be incorporated through scavenging by Mn(IV) oxide. Laboratory evidence shows that some other bacteria on nodules can reduce the Mn(IV) oxide to Mn2+ in the presence of an appropriate source of reducing power in the form of organic carbon. The reduction of Mn(IV) oxide mobilizes manganese and trace metals such as Cu, Ni, and Co in the nodules. These bacteria reduce Mn(IV) whether oxygen is present or not. They could be harnessed in the bioleaching of nodules. The manganese(IV)-reducing bacteria and the bacteria that neither oxidize nor reduce manganese are thought to keep the organic matter in the nodule environment sufficiently low to favor bacterial manganese-oxidation, which in laboratory tests is inhibited by excess organic carbon. The preying by the protozoa on the different kinds of bacteria is thought to control the bacterial population on nodules and ensure continued growth of the manganese-oxidizing bacteria and the nodules. To eliminate interference from seawater salts in the bioleaching of Mn, Cu, Ni and Co, it may be more advantageous to use terrestrial Mn(IV)-reducing bacteria than the Mn(IV)-reducing bacteria resident on nodules. The terrestrial bacteria would allow bioleaching in fresh water. Unlike the marine bacteria, most terrestrial bacteria studied to date reduce Mn(IV) oxide only anaerobically. A general outline for an anaerobic bioleach process of marine nodules with freshwater bacteria is discussed.

Jan 1, 2000

By Ted D. Maki, Joseph B. Taylor

INTRODUCTION FMC&apos;s Paradise Peak mine, located 13 Ian (8 miles) south of Gabbs, Nevada, became opera¬tional in April of 1986 (Figure 1). It was designed and built by Davy McKee, who was instrumental in equipment design and selection. It is the 7th largest gold deposit in the United States with defined reserves of 10.9 mt (12 million st) containing 34 million gms (1.1 million tr oz) of recoverable gold and 933 million gms (30 million tr oz) of recoverable silver. The mine extracts ore by the open-pit method, taking advantage of a 1.5:1 stripping ratio. The mill operates at a 3600 mt/day (4000 st/d) capacity. Crushing is done in three stages to achieve an ore size of minus 0.635 cm (0.25 inches). Grinding further reduces the ore to 85% minus 100 mesh. The ground ore is treated with cyanide in agitated leach tanks and then washed in the counter-current decantation (OCD) thickeners. Zinc is added to the clarified, deaerated pregnant solution to precipitate the precious metals. The precipitate is acid digested to eliminate excess zinc, filtered and retorted to drive off contained mercury. The retorted precipitate is then fluxed, melted and poured through cascade molds. Dore bars are cleaned for shipment and the slag is sent to an in-house slag treatment system. ELECTROSTATIC SEPARATION The dry electrostatic slag treatment system at Paradise Peak is the first installation of its kind. Electrostatic separation has been widely used in mineral processing since the early 1950&apos;s. A brief discussion of the theory behind the process is helpful to those not familiar to electrostatic separation. Charging and sepa¬rating slags, dry minerals or other materials by ion bombardment is the most common form of electrostatic separation. Millions of tonnes of minerals are processed each year by this method. In an ion bombardment separation, granular material is fed onto a grounded metal cylinder (or roll) and charged by a corona-producing electrode placed above the roll&apos;s surface (Figure 2). While both conductors and non¬conductors become charged, only the conductor is able to lose its charge. The charged non¬conductor, as it rests on the roll&apos;s surface, "sees" an oppositely charged image of itself in the metal surface. It is attracted to the image charge, becomes electrostatically pinned to, and moves with the roll&apos;s surface. The conductive particle also sees an image and is attracted to it. But upon touching the roll&apos;s surface, it discharges rapidly to the grounded surface and is thrown free from the roll&apos;s surface with a projectile motion. In the case of precious metal slags, the con¬ductive particles would be metallic prills of dore metal; and the nonconductive particles would be the slag, free of metal. Middling grains, generally, are in the form of a metallic prill encased in or incompletely liberated from slag. LABORATORY TESTING A number of precious metal slag samples have been tested in the laboratory. It has been found that the composition of slags vary widely from one refinery operation to another. Typical laboratory procedure is to crush and size the slag followed by two-stage lab-scale electro¬static separation. The conductor fractions from the first and second pass are then combined as a prill concentrate; middlings fraction and clean slag tailings from the second pass are held separate. Selected results from laboratory testing are shown in Table 1. It is important to note in Table 1 that the assays represent overall silver and gold, not metallic values. From experience, the laboratory results are typically lower in grade and recovery than the industrial installations. This is largely due to the hydroscopic nature of precious metal slags coupled with the high local humidity in Carpco&apos;s Jacksonville, Florida, location. Relative humidity in Jacksonville can range form 60-95%, while most mining locations in the

Jan 1, 1987

By M. Ge, C. Mborah

"The acoustic emission/microseismic monitoring (AE/MS) technique has widely been used in the mining environment for the monitoring of rock mass stability. A successful implementation of the method is hinged on being able to accurately determine the P- or S-wave arrival which is then used in the determination of the source of an event. Determination of the arrival phase is usually performed visually by a human expert. However, the mining environment is characterized by high noise levels and different noise types of varying intensities. Therefore, performing manual phase picking on time series data acquired from the AE/MS monitoring system can sometimes be tedious and time consuming if not impossible. In this paper, a proposal for improving manual P-wave phase arrival picks is discussed. The method involves the filtering of the AE/MS data using a stationary discrete wavelet transform (SDWT) method. After filtering, appropriate scales/levels of the multi-level decomposition process are then selected for further processing. The first step in the picking process after filtering, requires the computation of the signal power and the root mean square (RMS) using the signal amplitudes at each scale. The preliminary P-phase arrival time is picked using the results of the calculated power and RMS values respectively. The calculated RMS values are then superimposed on the calculated power values to determine the final arrival picks. If the arrival time picks obtained by both methods coincide or are within 5 to 10 sample points, the pick is considered reliable. The accuracy of the method was verified using AE/MS data from two datasets obtained from two separate underground mines. The results of the study showed that the picking method is reliable and robust. Based on the data studied, it can be concluded that the method will help reduce the time usually spent in processing AE/MS data from mines. INTRODUCTION A successful identification and location of the source of acoustic emission/microseismic (AE/MS) events is mainly dependent on the onset time of the P-wave phase. In many seismic networks, initial picks are performed by picking algorithms followed by manual picks by human experts. If the waveform signal-to-noise ratio (SNR) is low, the P-wave onset time picks by the human expert is generally reliable and accurate [1]. However, for waveforms with high SNR, picking of the P-wave onset could be a major challenge. Also, due to the volume of data acquired daily, manual processing could be stressful and time consuming [2]. These challenges continue to provide motivation for alternative means of providing reliable and efficient ways of improving manual phase picking. The AE/MS technique has been an intergral part of safety monitoring in mines [3] due to its capacity to identify zones of instability in a monitoring area. To obtain an accurate source of an AE/MS event however depends on the quality of the data used for P-wave phase picking [4, 5]. AE/MS data from mines are generally contaminated with excessive background noise generated by the different mining operations. The AE/MS signals of interest are mostly buried in background noise which makes it difficult to pick the P-wave phase arrivals [4]. Processing AE/MS data acquired in noisy environment remains a major challenge due to the weak nature of AE/MS signals [6]. As noted by [4] most of the data recorded in the mine environment are mainly due to events not related to AE/MS events of interest. The complexity of data in this type of environment makes the detection and picking of P-wave arrival by human experts a very difficult task."

Jan 1, 2017

By John R. Buzek

Thickeners are simply large tanks, usually circular in shape, which are designed to allow settling of solids and to operate with continuous overflow of clear water and underflow of thick pulp. The dimensions of the thickener are usually such that dewatering of very fine pulps may be accomplished while still 1 overflowing clear water. Some thickeners such as the Genter, Hardinge and Hydrotator types combine thickening with filtering2 while others such as the Dorr type are pure thickenersl,2,3. A schematic diagram of the traction type Dorr thickener is shown in Figure 1. This type is appropriate for tanks over 15 m (50 ft) in diameter. This tank employs a rotating raking mechanism which moves the settled material toward the central discharge. For this and other thickeners, it is common to have a shallow conical tank bottom in order to conform to the geometry of the raking mechanism, and it is often desirable to elevate the tank for ease of access to the product. When the tank does not rest on the ground, the forces resisted by the cone are large, and thick plates must be employed. The fabrication of these tanks requires temporary supports in order to erect the cone. A flat-bottomed tank could also be used because the settled material would soon form a hardened conical surface. However, because of the presence of large bending forces, even larger plate thicknesses are required than for the conical bottom. Significant cost savings often can be realized by the use of a scalloped-bottomed tank such as the one partially shown in Figure 2. The particular scalloped bottom discussed in this paper (and shown in the Figure) employs radial beams forming the outline of the cone-shaped bottom. These beams are supported directly by columns. Cone segments are hung between adjacent beams. The exact geometry of these segments depends upon the cone angle and radius, the number of radial beams, and their inclination. The geometry is chosen to satisfy certain clearance restraints and to utilize the material as efficiently as possible. The resulting thickness of these segments is significantly less than that of the

Jan 1, 1977

By D. A. Wheeler

Invented by Pierre Boutin in the early 1960s the column was a complete break from the conventional flotation cell. 1. When used as a rougher-scavenger, the column is excellent. 2. When used as a cleaner, the results can be spectacular. The very first column tests were carried out using it as a rougher-scavenger in a reverse float where silica was floated from iron. It produced a concentrate underflow as good or better than that produced from roughing and scavenging stages in cells, The froth tailing overflow was better than that produced by several stages of cell cleaning. Column scaleup progressed rapidly from the two inch diameter unit to a semi automated 12 inch diameter column on material from the Iron Ore Company of Canada. A change in operating philosophy within IOCC brought all flotation development, column or conventional, to a halt. At that time, IOCC had an exclusive right to the use of the column in Canada in the field of iron ore. We moved into the field of sulphides. A Canadian copper producer sent ore for testing and the results led to their purchase of the first commercial size of column - a 36 inch diameter machine. It was a mechanical disaster. It took several years to raise sufficient funds to return to that mill with our basic 18 inch square unit. It was to be tested and modified in order to learn how to properly design a large column. Originally used as a rougher- scavenger, it had to produce tails equal to the final tailings from this well run plant. It did, and did so while producing a rougher-scavenger concentrate almost equal to the plant final concentrate. It was finally used as a cleaner and produced concentrate 5% higher in copper than the plant final concentrate with equal cleaner tails. This phase of the column development was carried out under very difficult circumstances. The mill superintendent had realized that if he took the froth removal system from our original mechanical disaster and applied it to a conditioner while injecting air, he would have a flotation cell. He had personally applied for patents on this Maxwell Cell. Our development work was done in his mill and as our results became better and better, our difficulties became worse and worse. We finally had to terminate this work. The 18 inch column installed at Mines Gaspé 14 years later was identical to the one removed from Opemiska. The 18 inch column was tried on various ores over the following years and always produced excellent results. However, it was not really a production size unit. We had always aimed for the 72 inch column (72" x 72" x 44&apos; 9"). Prior to this huge machine, we needed the intermediate 36 inch column. The failure of the original 36 inch diameter unit at Opemiska had raised the possibility of short circuiting inside the column as the cross section increased. The first 36 inch square unit was tested in parallel with the proven 18 inch column. If the underflow of the 36 incher was not as good as that of the 18 incher, short circuiting was a possibility in the larger unit. It had been designed for insertion of drop in partitions, four feed points had been provided and the 36 inch column would have become a modular unit of four 18 inch columns. Testing showed the underflows of both columns to be identical. There was no short circuiting in the 36 inch column. Once we had the 36 incher, we had no fear of the 72 inch column. It is permanently partitioned into four 36 inch columns but uses only one set of instrumentation. All the component parts of the 72 inch column come from the 36 inch unit. In spite of our results, the mining community did not believe the column could work. Finally, in 1980, Mines Gasp6 ordered an 18 and 36 inch column for their byproduct molybdenum

Jan 1, 1988

By V. C. Rogers, K. K. Nielson

INTRODUCTION The potential health impact from uranium milling operations is mainly associated with long-term releases of radioactive contaminants from the mill tailings.(1,2) The major mechanisms for mitigating these potential releases focus on increasing the tailings containment with the addition of migration barriers such as thick earthen covers and clay liners. Some limited investigation has also focused on reducing the radionuclide source terms. This alternative approach has some desirable features, but stringent cost requirements are placed upon source removal methods in order for them to be economically favorable. A cost effectiveness evaluation is presented herein, in which costs for containment methods are used to establish maximum cost guidelines for the source removal methods. The 230U decay chain, shown in Figure 1 contains the radioisotopes of major concern, 230Th, 226Ra, 222Rn and its short-lived alpha-emitting daughters, When the uranium is removed in the milling process, these isotopes, which are normally not removed remain in the tailings solids. The general pathways to the environment from uranium mill tailings are depicted in Figure 2. The dominant pathways are radon to the atmosphere and radium and thorium to a groundwater aquifer. The magnitude of the contaminants released into these pathways can be characterized by the following parameters: 1. The 230Th and 226Ra content. 2. The emanating power for radon. The fraction of radon generated that is free to migrate. 3. The leach rates. Source reduction methods seek to reduce the magnitude of any or all of these parameters. In general the dose rates are proportional to the source terms. For radon production the source term is proportional to the product of the radium concentration, R, and the emanating power, E, so that the risk is reduced by the same factor that reduces RE. For example, if the radium content is reduced by 95 percent, but E is increased by a factor of two, then the dose rates are reduced by an order of magnitude. The source term for the groundwater pathway is not as easy to identify. The peak in the radium concentrations in the accessable water body can also

Jan 1, 1981

By Ernest E. Thurlow

Introduction Today, more than 50% of rail-carried commodities are mineral industry related, with coal being the most important single commodity moved by rail. In 1980, coal accounted for more than 5.7 million of the total 22.6 million carloads moved by rail. Metallic ores were third behind grain, with more than 1.4 million carloads. Crushed stone, gravel, sand, and other non-metallic minerals totaled almost 2.3 million carloads. Chemicals and allied products, including fertilizer and coke, added another 1.7 million cars, with petroleum and petroleum products totaling 300,000 carloads. Coal Several things happened in the 1970s that gave rise to increased consumption of coal-particularly western coal-and to its dominant position among rail-carried commodities. First, the Clean Air Act of 1970 required generating plants to make significant reductions in sulfur dioxide emissions. To comply, utilities could either invest in scrubbers or switch to low-sulfur western coal. Many opted for the latter. By 1972, increasing demands by the utility companies halted what had been a 25-year decline in national coal consumption. Second, the Arab Oil Embargo of 1973 put an end to cheap oil and gas, limiting their future as fuels for electric generating purposes and increasing the potential for coal. Third, in 1977, President Carter announced an energy program with coal as its cornerstone, calling for an annual two-thirds increase in national coal production by 1985. He also called for conversion to coal by utilities and large industrial users. Finally, he proposed a 10-year, $10 billion program to encourage domestic coal production and stimulate development of export markets. Coal bounded into world prominence. Foreign demand for steam and metallurgical coal increased tremendously, while US demand for western coal also shot up. This meant greater demands on the transportation sectors that traditionally carried coal to market. Many railroads began programs to serve the coal industry. One example is Burlington Northern&apos;s commitment to handle increased western coal tonnages. The company spent more than $1 billion in recent years to develop a system capable of moving more than 91 Mt (100 million st) of coal each year. Other leaders in this renewal were Norfolk and Western, Union Pacific, Santa Fe, and Southern Pacific railways. The importance of the rapid growth of coal traffic to the railroads is shown in the accompanying table, which gives percentages of total tonnages hauled and revenues attributed to coal. With coal providing the railroad industry with a substantial share of its revenues, there is keen competition among the rail companies themselves and among railroads and other transportation sectors for coal haulage. But there is also cooperation when more than one railroad is involved in delivering the coal from mine to market or when a combination of transportation modes is more economical. The latter is represented by Conrail&apos;s interest in working with the port authorities of New York and New Jersey to establish a new coal port that would serve not only export markets, but also utilities and industries in the northeast. Iron Ore Next to coal, iron ore (taconite) is the most important single mineral commodity handled by railroads. In Minnesota, where most iron ore is produced, rail transportation is primarily by Burlington Northern and the Duluth, Missabe, & Iron Range Railway Co. (DM&IR). The DM&IR, owned by US Steel, serves several producers on the Mesabi Iron Range. Two of the larger producers, Erie Mining Co. and Reserve Mining Co., also own railroads that operate between the mines and the ports of Silver Bay and Taconite Harbor on Lake Superior. Several iron mines and taconite plants in Michigan are served by the Chicago and Northwestern, and the Soo Line. Total 1981 shipments of taconite and iron ore are estimated at 55.9 Mt (55 million It), compared with about 61 Mt (60 million It) in 1979 the most ever shipped in one year. Still, with annual production capacity of the eight Mesabi Range

Jan 10, 1982

By Boleslav Taraba, Antonín Taufer

Thermovision monitoring of surface temperature has been examined for the purpose of detection of spontaneous combustion of early stage. Measurements were carried out in bituminous and subbituminous coal mines. It was found that surface temperature heterogenities can be considered as common phenomenon in mine openings. However, the most of the observed temperature heterogenities were not connected with coal spontaneous combustion of early stage. These, so called [non-spontaneous combustion] temperature heterogenities in surface of mine openings were found to be caused by: l. Different conditions of heat transfer between opening surface and air stream; 2. Low temperature (natural) oxidation of coal; 3. Dissipated heat loses of working machine. Sites of early stage spontaneous combustion of coal were characterized by: 1. More significant, local rise in surface temperature comparing with low temperature coal oxidation; 2. Evolution of carbon monoxide with local concentration near the surface of such site more than about 15 ppm (bituminous coals) or 30 ppm (subbituminous coals). INTRODUCTION Mine fires are a serious hazard to life and result in a considerable loss of coal reserves. Thus, great attention is paid to investigation of new methods enabling already early detection of self-heating of coal. The earlier a fire is observed, the easier would it be to deal with. Detection methods used in coal mines are mainly based on: 1. [Mine air analysis]. Changes in composition of mine air are examined. Products of spontaneous heating of coal, not being present in mine air under normal conditions, are considered as indicators of anomalous behaviour of coal (carbon monoxide, unsaturated hydrocarbons) (Mackenzie-Wood and Strang, 1990). 2. [Temperature measurements]. Sites of higher temperature level comparing surroundings are looking for underground using various direct or indirect methods of temperature measurements. This paper summarizes experience from thermovision monitoring, as one of the latest indirect temperature measurement technique, used for the purpose of early detection of self-heating of coal in mine openings. INFRARED THERMOVISION Infrared thermovision device of AGEMA Infrared systems AB (Sweden) was used for the in situ measurements. The constant value of emissivity factor, 0.95, was pre-set as the most probable level of emissivities of coal and rocks in mine conditions.

Jan 1, 1993

By R. Shields

"Hard rock mines move large volumes of material. But the conveyors essential to this movement suffer lots of problems, including carryback, spillage, dust, and belt wander. Many of these issues are created at the transfer points where the belts are loaded. This presentation will start with an overview of the latest transfer point systems that incorporate safety and serviceability by design. It will then report on recent projects. One project is the Conveyor B Transfer at the Coeur Rochester silver mine near Lovelock, Nevada. Here performance was improved with the installation of belt support cradles, engineered chute walls with external wear liner, and a compact dust collector. A second project is the improvement of belt support idlers and skirtboard sealing in a long loading zone of a conveyor under the concentrator at an Arizona copper mine. THE IMPORTANCE OF TRANSFER POINTS Conveyors are essential in mine material handling. Their performance will often make the difference in mine productivity and profitability. A key to conveyor performance are the transfer points where the conveyor belt is loaded and discharges. At these transfer points, many of the problems encountered with the conveyor system are created. The problems that arise at the transfer stations include off-center and segregated loading, mistracking, blockages and plugging. And there are a variety of problems originating with fugitive material—cargo released in the form of carryback, spillage, or airborne dust—that are created or exacerbated in the transfer points. These fugitive materials turn into other conveyor-related problems, including shortened component life, unplanned outages for conveyor maintenance, unexpectedly high labor costs for maintenance and cleanup, worker health and safety problems, and issues with community relations. As estimate attributed to the U.S. Mine Safety and Health Administration (MSHA) notes that 85 percent of conveyor maintenance is a result of fugitive material at conveyor transfer points. TECHNOLOGIES TO IMPROVE TRANSFER POINTS A number of techniques to improve the performance of conveyor transfer points have been developed. They include: • Absorbing impact, to prevent damage to belt, belt support systems, and conveyor structure, and reduce spillage caused by the impact energy arising from long material drops or large lumps of material loaded onto the receiving belt. • Stabilizing the belt line, to eliminate sag that allows spillage and encourages material entrapment that can damage the belt. • Protecting the skirtboard seals, by sheltering the sealing strips from the mass of cargo pushing material toward the edges of the belt. • Using self-adjusting, multiple-barrier edge seals, to keep material on the belt by providing long-lived seal efficiency with minimum maintenance requirements. • Controlling and centering the belt’s path, to eliminate belt wander and the spillage and damage it can cause. • Allowing dust to settle, through the incorporation of chutework that provides expanded stilling zone and the use of dust curtains at the points where the belt enters and exits the transfer. • Adding supplemental dust control systems, through the installation of systems for passive dust gathering, water-based dust suppression, and/or active dust collection."

Jan 1, 2017

By H. Staton, N. Michaelson

"INTRODUCTION Lundin Mining’s Eagle Operation, located in Marquette County, upper Michigan consists of a mine and concentrator that process a Ni-Cu-(PGE) ore. The mine itself is an underground operation and the ore is trucked from the Eagle mine site 66 miles to the Humboldt mill where both nickel and copper concentrates are produced. Operations began in 2014 and continued with an average production of approximately 23,000 tonnes of Ni and 20,000 tonnes of Cu per year for the first three years of operation. The Humboldt Mill, currently operated by Lundin Mining, processes the nickel/copper ore from Eagle Mine and performs the concentration through several stages of flotation. Originally built in 1954, the Humboldt Mill began operations as an Iron Ore processing facility, and was converted to a gold processing plant during the 1980s. Various groups owned the site after closure of the gold operation. Eagle purchased the Mill in 2008 and invested more than $280 million to clean up historic mine waste and refurbish the plant (see flowsheet in APPENDIX). This paper will overview the ramp-up of the operation, and discuss some specific areas that went well and some that could have gone better during the design phase. HIGH LEVEL DESIGN CRITERIA Coarse ore arrives from the mine and is stored in a building with a capacity of 10,000 tonnes. Ore is crushed in three stages to a final product size of 100% Passing 12mm. Crushed ore is conveyed to the mass flow fine ore bins, which have a live capacity of 3000 tonnes. Crushed ore is fed to two parallel 900 HP ball mills, charged with 3” high chrome grinding media, and ground, in closed circuit with hydrocyclones, to a p80 of 110 microns in the flotation feed. Soda ash is added to the mills based on process water hardness. Cyclone overflow passes to the bulk conditioner where Sodium Isopropyl Xanthate (SIPX) and Methyl Isobutyl Carbinol (MIBC) are added. The slurry is pumped to the rougher bank for bulk flotation. Rougher concentrate is fed to a 500HP vertically stirred regrind mill and is ground to a p80 of 50 microns, using 5/8” high chrome balls. Rougher tailings report to the tailings thickener. The regrind cyclone overflow is sent to the single stage cleaner bank followed by two scavenger banks. The cleaner concentrate is pumped to the copper-nickel separation circuit. Cleaner tailings report to the tailings thickener. In the copper-nickel separation circuit, the pH is raised to 11.5 and a copper concentrate is recovered from the bulk cleaner concentrate in two stages, namely the copper roughers and the copper cleaner columns. The concentrate from the copper flotation columns is the final copper concentrate. The tailings from the columns return to the head of the copper-nickel separation. The nickel concentrate is the copper rougher tailings stream."

Jan 1, 2018

By Brenda J. Wolfe

Following successful laboratory and pilot plant investigation of dewatering characteristics of molybdenum concentrate produced by a Chilean copper/moly producer, two (2) DENVER SALA 500 Series Tube Press Filters were installed in 1994 to handle the total production volume. Moisture requirements in the filter cake have been consistently met, eliminating the need for thermal drying after filtration. Operating data and Tube Press Filter are described herein.

Jan 1, 1995

By R. M. Williams

Introduction Ring roller mills are efficient size reduction machines embodying the following machine functions: 1) Size Reduction: from a feed size of 2 in. to various size products as fine as 400 mesh, or granulation from 2 in. to 6 mesh containing a minimum of fines. 2) Drying: fluid bed drying of wet feed materials. 3) Conveying: negative pressure air system results in dust-free conveying. 4) Beneficiation: liberation and removal of foreign materials from the auxiliary bottom discharge simultaneously while grinding. Operation Operation of the roller mill system, Figs. 78 and 79, is completely automatic. A feeder introduces raw material into the grinding chamber in the base of the roller mill at a rate determined by pressure variations within the roller mill as material is ground and air conveyed away. The feeder will be activated by the reduced static pressure and introduce unground material into the system, thereby maintaining a constant differential pressure. Plows located ahead of each grinding roll direct the material upward and between the grinding rolls and the bull ring where it is ground to size. The grinding rolls are free to swing out centrifugally, which forces them to bear upon and grind the raw material. Being free to pivot axially, the rolls automatically assume the proper position for grinding and need no adjustment to compensate for wear. Rotation of the rolls is by rolling contact. Once ground to size, material is swept out of the mill by the controlled air flow from the centrifugal main mill fan. Air suspended, the ground material passing through the adjustable spinner separator. In the separator, an accurate size classification is made which allows product-size material to be conveyed away from the roller mill while oversize material is returned internally for additional grinding. By varying the speed of the separator drive, more or less cyclonic action is achieved at this location, effectively controlling product size. Product sized material is removed from the airstream by the cyclone collector. Typically, a conveyor or product storage bin will be located to accept the product as it exits from the bottom of the cyclone collector.

Jan 1, 1985

By Santhanam Sundaram

The potential for the utilization solar energy for the production of industrial process steer in mineral processing industries is discussed briefly. To advance the technical and economic feasibility of solar-thermal systems applications, U.S. Department of Energy has funded several projects. NL industries of Newberry Springs, California is one among them and has been chosen as a case study. The NL Industries solar system consists of 950 square meters of Del, single-axis, tracking, parabolic-trough, concentrating collectors. It was designee to produce a portion of the 420°. 530°k. steam utilized in a trying operation reduce the moisture content of hectorite ore from 10 percent to 4 percent. (Hectorite is an hydrous magnesium silicate which, when refined, is of significant commercial interest because of its applications in various chemical and food processes.) is estimated that implementation or this solar system could result an annual savings of 3.545 billion KJ ,3.360 billion BTUs), or the equivalent of 90 cubic meters (600 barrels) of oil and a net reduction of 1,860 kilograms (4100 pounds) of air pollutants annually. The technical, economic and institutional issues encountered in the course of this project are also discussed. Finally, the impact of the commercialization of solar energy applications in mineral processing industries is evaluated.

Jan 1, 1981

By McGuirk. C. H.

The Port of Conneaut, long regarded as a leader in advanced bulk material handling technology, recently incorporated into its existing facility a new, multi-million dollar coal transfer-storage system. While the basic principles of conveyorized bulk material handling remain unchanged in the new system, design modifications such as rotary dumping of railroad cars, automated car feed to dumper, belt rip detection for conveyor belting, bucket wheel continuous production, and other innovations make The Pittsburgh and Conneaut Dock Company&apos;s complex worthy of classification as the "Port of Tomorrow - Today". The Pittsburgh & Conneaut Dock Company, the "Port of Tomorrow - Today", 1s a bulk materials storage and transfer facility located on Lake Erie, 68 miles east of Cleveland, 30 miles west of Erie, PA., and 138 miles north of Pittsburgh. Figure 1 shows an aerial view of the facility. It is designed to handle over 25 million tons per year of metallurgical and steam coals, direct shipping ore, fine ores, concentrates, pellets, limestone, and dolomite. Inbound shipments of iron ore and limestone are received from Minnesota, Michigan, Labrador, and Quebec. Coal arrives from Pennsylvania, Ohio, Virginia, West Virginia, and Maryland. Rail service is provided by the Bessemer & Lake Erie Railroad Company.

Jan 1, 1980

A combination field and office study of mining districts in Southwestern U.S.A., conducted in 1949 - 1950 by United Geophysical Company, embraced all of the essentials for a successful exploration project including factors such as geology water, transportation and many more. Perhaps the best known of several high priority areas chosen from the study was the Twin Buttes or Pima District in Arizona, however, the Turquoise District also ranked very high on this priority schedule. Following the discovery of the Pima Mine in 1951 and 1952, United initiated additional preliminary geological, geophysical, drilling and claim location work in the Turquoise District, but stopped shortly thereafter, when ASARCO began adjoining competetive activity. During the summer of 1957, sometime after ASARCO abandoned their claims in the area, Minerals Exploration Company began intensive exploration in the Courtland-Gleeson part of the District. This program was terminated in 1958. Later Bear Creek Mining Company supplemented the discovery endeavor with their own program during the period 1960 through 1962. This paper summarizes most of the geophysical work done by all of the groups except ASARCO and is intended to complement the recent paper "General Geology and Some Structural Features of the Courtland-Gleeson Area, Cochise County, Arizona" by Mr. Otis McRae of Bear Creek Mining Company.

Jan 1, 1966

By A. O. Drysdale

From a mining point of view the Portland Cement Industry falls logically into the Industrial Minerals Division. Portland Cement manufacture requires four principal materials, lime, silica, alumina and iron. The usual source of lime is limestone. Silica sources include clay, shale, quartz, sandstone and sand. Alumina is usually provided by selecting a silica source with a suitable alumina content such as clay and shale. "Cement Rock" is occasionally found in which all four ingredients are found in natural combination or sufficiently close that only minor adjustments are required. In the absence of iron oxides a white cement can be produced. All raw materials used in Canada contain some iron and in most plants further minor quantities of iron are added to promote fluxing, hence the typical grey colour of Portland cement. The modern period in cement manufacturing started in 1824 when Joseph Aspdin of England patented a process for the manufacture of a modified hydraulic cement which he called "Portland Cement". Hydraulic cements are cements that harden under water. Although natural cements, made by calcining and grinding natural argillaceous limestone were produced in Canada as early as the 1830&apos;s, Portland cement was not produced in Canada until 1890, about 18 years after it was first made in the United States.

Jan 1, 1964