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By A. D. Hammond

"INTRODUCTION The recent drop in iron ore prices has claimed an unwitting casualty in vanadium. Vanadium is dominantly produced as a co-product of steel production. The recent shutdown at EVRAZ’s Mapochs steel-vanadium operations in South Africa has cut-off 10% of the world’s vanadium output. Earlier in February, 2015, Australia’s Atlantic Ltd. suspended operations at its Windimurra iron-vanadium complex after a fire caused severe damage at the plant. Since then, challenging market conditions has resulted in the permanent closure of this operation. This decoupling of vanadium production from integrated iron ore producers presents an opportunity for vanadium exploration companies to target single-product deposits. This trend is manifested in the increasing number of pure-play vanadium ventures around the world.The mine production of vanadium is dominated by three countries namely, China, South Africa and Russia, which collectively produce 98% of the world’s vanadium. While the use of vanadium as an alloy in the steel industry remains by far the largest application for this metal, growth in demand is expected to come from the renewable energy industry in the form of batteries for energy storage. This paper summarizes the current state of vanadium exploration, briefly describes the vanadium market and lists sources of future demand for the metal. TREND IN VANADIUM EXPLORATION The most active participants in the current vanadium exploration are pure-play vanadium ventures in Brazil, South Africa and Australia. These ventures are targeting deposits with higher vanadium grade than those typically mined in titaniferous magnetite operated by integrated steel-vanadium producers. The 10% shortfall in vanadium production resulting from the shutdown at EVRAZ’s Mapochs is expected to be partially compensated by Largo Resources’ Maracás Menchen vanadium operation located in the State of Bahia in Brazil, which commenced operation in August 2014 and is expected to reach design capacity of 11,400 tonnes of vanadium pentaoxide (V2O5) by the end of 2015. The estimated production1 for 2015 is 6,500 tonnes of V2O5 at a target production cost of $3.60/lb. This is the first vanadium operation to be placed in production in the Western Hemisphere in over 35 years."

Jan 1, 2016

By Ireneusz Tomza, Jolanta Lebecka

INTRODUCTION Radioactive deposits were observed in 1972 in some of the Upper Silesian coal mines. They were located mainly in the drains in galeries and on the inside surfaces of water pipes. They also caused some problems by accumulating in water pumps. It has been postulated that the deposits are produced by natural radioactive waters seeping from the rocks. Investigations were initiated to answer the following questions: - What is the composition and the amount of radioactivity in the deposits? - What radioisotopes are present in the water? - How are the radioactive deposits formed? - Do the radioactive waters also occur in other mines? - How does the radioactivity of the water depend on chemical composition? - What is the origin of the radioactive water? - Does the water and the deposits cause radiation hazards for miners? -How can the radiation hazard be reduced? METHODS OF MEASUREMENT Determination of Radium Isotopes in Water The commonly used methods of radium determination in water are either based on measurements of the radioactivity of 222Rn which is in equilibrium with 226Ra, or on the detection of alpha particles of the radium radioisotopes after chemical separation of radium from the water sample. The method based on radon activity measurements is very sensitive and does not require any chemical Separation, but it can be used for determination of 226Ra from the uranium series only, because the thorium daughter 220Rn has too short a half-life (55s to yield the required accuracy. The method developed by Goldin, 1961 [2] involved alpha-particle measurements in thin layers of RaS04 and BaS04 separated from the water. This method is not convenient for saline water and water with high barium concentration because the amount of barium carrier in this case is too large to obtain a thin layer of precipitate with sufficient activity. The Upper Silesian carboniferous waters are often saline with high barium content, so the method described by Goldin was not convenient for this case and it was necessary to change the detection system and modify the chemical preparation. The procedure developed by the authors for the determination of radium isotopes in water was as follows: - Depending on the Ba2+ content and the required sensitivity of measurement, a water sample of 200 cm3 to 3 dm3 was taken. - 10 cm3 of 0.25 M citric acid and 5 cm3 15M ammonia was added to form complex Ba2+ ions and avoid the immediate precipitation of BaSO4. (This was repeated as long as the addition of BaC12 did not form a precipitate.) - 1 cm3 of 1N solution of Pb(N03)2 as a carrier for radioactive isotopes of lead and 10 cm3 of 0.1 N BaC12 as a carrier for radium were added. - The sample was heated to the boiling point and the precipitation of RaS04, BaSO4 and PbS04 with 50% H2SO4 was carried out. - After several hours the sample was centrifuged and the precipitate was purified by washing with nitric acid and distilled water. - The precipitate was then redisolved in 20 cm3 0.125 M Na2EDTA and 3 cm3 6M ammonia and reprecipitated from the solution by dropwise addition of acetic acid to d pH of 4.5. At this value of pH, precipitation occurs only for the barium and radium sulfates, while lead and all other radioactive elements remain in the solution. The date and time of deposit precipitation was recorded. - The final barium-radium sulfate mixture was washed with distilled water and transferred to standard measurement vials. - Each vial containing a deposit had 6 cm3 of distilled water added and was then shaken vigorously. 12 cm3 of liquid gelling scintillator (INSTA-GEL UNISOLV-1 type) was then added and the vi 1 was shaken again. After a while the scintillator turns into a milky gel in which the deposit is uniformly distributed. - The standard sample of 226Ra was prepared in the same way. - The activity of the samples was measured using a liquid scintillation spectrometer. (In this case the TRICARB 3320 produced by Packard Instruments, was used). Tests run on standard radium solutions provided by Amersham Radiochemical Centre indicated that this method of measurement enables one to achieve an efficiency of almost 100% (within measurement error). For alpha particles no quenching effect was observed for the BaS04 concentration in the range up to 80 mg of BaS04 per 1 cm3 of liquid scintillator coctail (Fig. 1). This provides a sensitive determination of radium in water with high barium content and also in saline water. In saline water the solubility of barium sulfate is much higher than in

Jan 1, 1981

By R. J. Hamilton

Considerable success has been achieved in recent years in reducing concentrations of respirable dust in UK coalmines; thankfully this has been reflected in significant improvements in pneumoconiosis statistics. However, continuous increases in mechanisation and in rates of working, accompanied by more stringent dust regulations, necessitate the introduction of improved and often more sophisticated methods of dust control. The increased application of exhaust ventilation and air cleaning techniques seems likely to be the most important step forward in dust control in coalmining during the next few years. A typical drivage in a UK mine is not more than 16 ft (5 m) wide and is supported by steel arches. In the past the majority of such roadways were driven by the conventional means of blasting and loading. Since the main purpose of ventilation was to clear blasting fumes and to provide adequate scouring of firedamp, not surprisingly it was usual to have 'forcing' ventilation. Present systems of operation are making increasing use of heading machines and it is not surprising that well over half of all drivages are now ventilated by exhausting methods. (1) The great majority of coal getting in the UK is from longwall faces, usually in seams of less than 6 ft (1.8 m) thickness. Although this method of min¬ing has many advantages in terms of safety and productivity, dust control is made more difficult because of the need for men and machines to work in the same ventilation stream. Much emphasis has been placed on improving the design and methods of opera¬tion of longwall machines to reduce dust production, and improved ventilation has provided greater dilution, and in consequence, reductions in dust concentration. On all machines water is now fed through internal galleries to the cutting elements. Unfortunately, one of the difficulties associated with mining thin seams is that there is some unavoidable cutting in the roof or floor and in many seams there are faults to be traversed and dirt bands which must be removed with the coal. Under these conditions it appears that the use of exhaust ventilation and air cleaning offers the most promising way of achieving satisfactory dust conditions.

Jan 1, 1980

By T. R. Ley

"INTRODUCTION Brutally honest mining executives will tell you that “safety is job #1 but.. dollars are also extremely important”. When economic times are tough, as they are now in mining, it can feel like you need to choose between safety and cost control. Leading edge companies know that it is excellence at both that provides the best performance. FLUID POWER SYSTEMS Fluid power systems are ubiquitous in mining – longwall systems, continuous miners, drills, haul trucks and excavators, etc. And fluid power systems are getting more powerful and using higher pressures to move greater volumes and achieve higher productivity. Average operating pressures in mobile equipment has been rising to 450 bar (6,500 psi) over the last decade. With increased system pressure comes ever increasing levels of stored energy. Stored energy that, if released in an uncontrolled fashion, can do tremendous harm to people and things.Fluid power systems are composed of pipe, flexible hoses, pumps, valves and actuators. They are filled with fluid (mineral oil or fire-resistant) to transmit power to where it’s needed for performing work. With time and use fittings can loosen, component performance diminishes and hoses wear and tear. In hostile environments, like coal mining, hydraulic hoses can fail in as little as 2 years according to DGUV (umbrella association of German Accident Insurers), maybe sooner in some particularly challenging environments. FLUID POWER SYSTEM SAFETY While appearing simple and benign on the surface, fluid power systems, particularly such systems on mobile equipment, have many potential hazards. Hazards that most workers aren’t aware of – particularly young, new hires. And those hazards that can quickly turn deadly. During the past couple of years, including at last year’s SME Conference, I talked to a wide variety of mining people about high pressure fluid injection injuries. Some people, often with little or no field experience, never heard of it. But, people with mine operations / maintenance experience were almost always aware of such injuries.Some had even seen it happen right in front of them. But most claimed it didn’t happen all that often. It just wasn’t seen as a problem!"

Jan 1, 2016

By François Martin, Roland Plassart, François Laigle

"MOTIVATION The storage of nuclear waste gives rise to a lot of innovative issues for underground space technologies. The nuclear activity of waste generates temperature effects on both the lining and the surrounding rock mass. In order to avoid these unfavorable effects, the number of waste containers is controlled in the gallery section. The galleries section has consequently ranged from a diameter of 4 to 6 meters. However, in the case of intermediate-level nuclear waste, due to the lower temperature, the number of containers in a cross section of a gallery is assumed to be safely increased. Despite technical difficulties, a large diameter gallery (typically 12m excavated) seems to be more cost effective. This paper demonstrate the feasibility of such a gallery in the specific context of CIGEO (French Geological Repository Project for nuclear waste), located in the Bure area in the north-eastern of France. The geology of Bure at the project’s depth (about 520 meters) consists in a quite homogeneous layer of Callovo-Oxfordian argillites (COx argillites). With such an overburden, this rock is known to have a high plastic behavior with significant time-dependent effects. These mechanical properties require complex behavior laws in computational modeling used for the design of underground facilities.EXPERIENCE OF LARGE DIAMETER GALLERIES In order to identify some specificities in the design of such openings, a research has been run on existing galleries of large diameter (more than 10 meters), like the tunnel of Chamoise, Lötschberg, Tartaiguille or Saint-Martin-La-Porte. Two of these representative examples are detailed hereafter. Tunnel of Chamoise (French Jura) In spite of a lighter overburden (400m versus 520m), an anisotropic initial stress state and a variable rate of carbonates, the highway tunnel of Chamoise is quite similar to the planned gallery for intermediate-level nuclear waste, with a geology very close to the CIGEO’s one in some sections. Because the northern tube encountered difficulties during the excavation a few years before (several over-excavations and spalling, especially in the bench), the excavation of the southern tube has been carried out in two steps, but with a high bench and a small curved invert (Figure 1). The design of the immediate support consists in long bolts and a shotcrete layer. The concrete lining is about 50cm thick."

Jan 1, 2016

By B. H. Johnston, B. N. Doyle

Introduction In 1983-84 Kidd Creek Mines Ltd. in Timmins, Ontario, started up the second commercial zinc pressure leaching plant based on the Sherritt Gordon Mines Ltd. direct pressure leach process. The first plant using this process was that of Cominco Ltd. at Trail, B.C. (Parker, 1981; Martin and Jankola, 1985). The Kidd Creek plant was constructed to take advantage of the extra capacity that the existing purification, cell house, and casting sections had in comparison to the roasters and acid plant. In addition, pressure leaching offered flexibility in other areas such as treatment of low grade concentrates. The process design package and the detailed design of the autoclave system was carried out by Sherritt. The detailed design for the rest of the plant was done by Cambrian Engineering Ltd. under the direction of Kidd Creek. The construction and commissioning of the plant has been described earlier (Johnston, 1983). This paper describes, briefly, the application of the direct pressure leaching process for zinc concentrates and the integration of the pressure leaching plant into the main zinc plant at Timmins. It discusses some of the startup problems and changes made to the plant. The damage following a fire inside the autoclave is also described. The subsequent repair and recommissioning of the autoclave has enabled Kidd Creek to demonstrate the originally sought advantages of pressure leaching and to take advantage of their extra cell house and casting capacity to produce more zinc. Process principles The development of the direct pressure sulphuric acid leach of zinc sulphide concentrate at temperatures above the melting point of sulphur and at elevated oxygen pressure has been described in detail in a number of papers (Martin and Jankola, 1985; Johnston, 1983; Veltman and Bolton, 1980; Kawulka et al., 1975; Veltman et al., 1977). The most significant aspect of the process is that sulphidic sulphur is converted to elemental sulphur in reactions that can be simply described as follows. ZnS + H2SO4 + 1/2 O2 ZnSO4 + S° + H2O (1) FeS + H2SO4 + 1/2 O2 FeSO4 + S° + H2O (2) The rates of these reactions increase with temperature. However, at temperatures above the melting point of sulphur, 12°C, the presence of a surface active agent is necessary to prevent the occlusion of the sulphide particles by elemental sulphur and the consequent termination of the leach. The zinc that is dissolved according to reaction (1), after purification of the zinc sulphate solution, is recovered by electrowinning. The iron that is put into solution according to reaction (2) is further oxidized to the ferric state and subsequently, as the acidity decreases, precipitates as hematite or as complex basic sulphates. MINERALS AND METALLURGICAL PROCESSING Examples of this are shown below. 2FeSO4 + H2SO4 +1/2O2 - Fe2(SO4)3 + H2O (3) Fe2(SO4)3 + 6H2O - Fe2O3.3H2O + 3H2SO4 (4) 3Fe2(SO4)3 + 14H2O - 2(H3O)Fe3(SO4)2(OH)6 + 5H2SO4 (5) (hydronium jarosite) The conventional zinc process involves concentrate roasting to produce zinc oxide and sulphur dioxide. The zinc oxide is leached in spent electrolyte and the sulphur dioxide is converted to sulphuric acid. The pressure leaching process can, therefore, be used to replace conventional concentrate roasting, sulphuric acid manufacture, and atmospheric leaching with one unit operation. Integration of pressure leaching into the Kidd Creek flowsheet A block flowsheet of the Kidd Creek zinc plant, with the new pressure leach plant integrated, is shown in Fig. 1. The original plant, which was constructed in 1972, is based on standard electrolytic zinc plant technology. Concentrates are dead roasted to remove sulphur. Sulphur dioxide is treated to produce sulphuric acid. Calcine is leached in two stages using cell-house electrolyte. Iron is precipitated as jarosite and solutions are purified to remove copper, cobalt, nickel, and cadmium to provide feed for a partially mechanized cell house. Zinc is plated from solution, and the resulting cathode is melted, cast, and sold (Johnston, 1983). The integration of the pressure leaching plant into the main zinc plant is relatively simple. As shown in Fig. 1, a fraction of the zinc concentrate and spent electrolyte are fed to the pressure leach autoclave, while part of the calcine from the roaster is used as a neutralizing agent in the oxidation step along with. the jarosite leach solution. The main exit stream from the pressure leaching plant is the slurry discharged from the oxidation step, consisting of neutral zinc sulphate solution, unreacted calcine, and precipitated iron oxides. This slurry is transferred to the neutral leach step in the main zinc plant. The leach residue from the pressure leaching step, consisting of basic iron sulphate and elemental sulphur, is washed and impounded along with the jarosite residue produced in the main plant.

Jan 1, 1987

By L. Cui

X. Huang (1997) recently suggested a poroelastic model for simulating the hydraulic fracturing breakdown pressure. His paper began with a discussion questioning Haimson and Fairhurst's (H&F) model. He claimed that the H&F model failed to lead to the Hubbert and Willis (H&W) model as[ a --> 0.] Huang also tried to explain why the H&F model could only work for special types of rock conditions. He pointed out that one possible reason could be that the Terzaghi's effective-stress concept had been adopted. The H&F model (Haimson and Fairhurst, 1969) was derived under the conditions that the borehole wall is fully penetrated (pp = pi) and a drained state is realized (steady pore-pressure field). Cui et al. (1997a) demonstrated that, under drained conditions, the total stresses in the penetrating poroelastic model (identical to the H&F model) degenerate into their counterparts in the elastic model (identical to the H&W model) as [a -- 0.] However, the effective-stress conditions are different for both of these models, because different pore-pressure conditions at the borehole wall were adopted. In the H&W model, p = po was assumed, i.e., the pore pressure field is not disturbed; but pp = pi was assumed in the H&F model. Assuming that Terzaghi's effective-stress controls tensile failure (that was the hypothesis adopted in both the H&F and the H&W models), only the following two special drained poroelastic cases may degenerate into the H&W model for very small a: • when the pore pressure at the borehole wall remains at the same level as the virgin pore pressure for a penetrating model and when the borehole wall is simply impermeable (i.e., the nonpenetrating model, Cui et al., 1997b). Therefore, simply setting a = 0 in the H&F model generally does not lead to the same problem described by the H&W model. On the other hand, when the porepressure boundary conditions do not correspond to the ones in the H&W model, a degeneration of the poroelastic model to the H&W model as [a – 0] should be questionable. The pore pressure at the borehole wall is generally dependent on the injection-fluid pressure and the penetrating conditions at the borehole wall, such as the existence of a filter-cake. For an impermeable wall, pp is independent of Pi, and it is basically an unknown [(how¬ever, pp -- po as t --oo).] For a fully permeable wall, pp is the same as Pi. Between these two extremes, pp should be a function of p; and the permeable condition of the borehole wall, which may be dependent of the leak-off coefficient cf. (the range of cf is from 0 to 1). Theoretically, for a rock with low permeability, a penetrating borehole wall is still possible. For saturated porous materials with very low values of a, poroelasticity shows that a pore pressure will be built up due to the stress concentration subjected to a nonhydrostatic in situ stress field (Cheng et al., 1993). This phenomenon is known as the Skempton effect. The variation of the pore pressure may be evaluated by [AP = 3 B(A rr + 06ee + A (Y,,)] (1) where B is the Skempton pore pressure coefficient. This pore pressure variation dissipates as time increases (it is totally gone as the drained state is approached). The rate of the dissipation mainly rely on the permeability of the formation. The dissipation is very slow for tight formations because their permeability is very low, and it is fast for rocks of high permeability. According to our analyses, the time period for this process could be from seconds (for sandstone's) to a couple of days (for stales). One possible reason that the H&C model did not agree well with the experimental results for rocks with low permeability might be that the time interval between the application of the loading and the fluid injection had not been long enough for the dissipation of the Skempton effect. The effective-stress law basically defines how much pore pressure contributes to the total stress. The difference between Terzaghi's effective stress and Biot's effective stress is that 100% of the pore pressure contributes to the total stress in Terzaghi's definition, while only a certain portion of the pore pressure ((ap) goes to the total stress in Biot's definition. Therefore, when the pore pressure at the boretole wall (pp) is determined according to Biot's effective-stress law, app in the total tangential stress is attributed to the pore pressure. In other words, Terzaghi's effective tangential stress is expressed

Jan 1, 1999

By William J. McCloy, Robert L. Gamble

A coal-fired fluidized bed steam generator rated at 12.6 kg/s has demonstrated the industrial and institutional application of fluidized bed combustion by burning high sulfur coal in an environmentally acceptable manner within a populated area. Designed to provide steam for heating and cooling, the unit was placed into service at Georgetown University in Washington, DC, and has already logged 5275 hours.* The unit has consistently operated well within Washington's strict regulations governing emissions of sulfur dioxide (SO2), nitrogen oxides (NOx), and particulates. The Georgetown program, cofunded by the university and the US Department of Energy, has shown that fluidized bed combustion is a promising way to realize the full potential of coal, the nation's most abundant energy resource. In a fluidized bed steam generator, a layer of granular limestone is placed on top of a grid designed to prevent particles from falling through. Air is passed upward through the bed with sufficient force to lift and float the particles. At this point the bed material resembles a boiling liquid. Crushed coal is then introduced into the bed and ignited. When high sulfur coal is burned, limestone absorbs the SO2 produced. The operating temperature of a fluidized bed steam generator is relatively low (approximately 871°C). As a result, NO, formation is minimal. Another potential problem encountered in burning coal is fly ash, a form of particulate. A feature of the fluidized bed design, developed by Foster Wheeler Energy Corp., is a fly ash reinjection system. Fly ash is reintroduced into the bed where it is burned again, thereby enhancing fuel efficiency. Nearly all remaining fly ash is captured by a baghouse dust collector. At Georgetown, the fluidized bed steam generator and all auxiliary equipment are enclosed in a structure aesthetically compatible with the surrounding university complex. The steam generator has two separate beds that can be fired independently, and operation to date has been with one and two beds in service, based on steam load demand. During several operating periods, when the fluidized bed unit was the only boiler in service, it supplied all the steam required by the 18.5 hm2 of buildings on campus. Initially fired with low sulfur coal, the unit was quickly switched over to higher sulfur fuel which will be used most of the time.

Jan 3, 1981

By Byron C. Hardinge, Michael Sokaski, Paul F. Sands, W. Loring III Mc Morris, David J. Akers

INTRODUCTION In the early days of coarse-coal cleaning by the dense-medium process in Europe, loess was one of the materials used for medium solids. Newly developed cyclones were used to reclaim and thicken the loess suspension. On occasion, when the cyclone thickener at the Maurits mine in the Netherlands plugged, the overflow was found to be filled with clean coal free of impurity. From this observation stemmed the concept of the cyclone as a cleaning device. Development took place during World War II in the laboratory of the Dutch State Mines under the direction of M. G. Driessen. A 15-tph (13.6 t/h) pilot plant was in operation in the laboratory by 1945, when, in September, Driessen made the first public disclosure of the process in a paper presented to the Institute of Fuel in London.51 In April 1945, operation of this pilot plant was observed by H. F. Yancey, who was so impressed with the potential of the cyclone process that, on his return to the United States, a research program on the cyclone was immediately started by the US Bureau of Mines (USBM). This work led to an AIME paper in 1946 in which the first details of cyclone operation on American coals were made available.52 Commercial adoption of the cyclone proceeded in Europe, initially in the Netherlands, Germany, and France. In the United States, where coal is less costly and most of the coals are easy to clean and respond well to simpler cleaning processes, the first cyclone plant was not built until 1961.53 THEORY Before discussing the theory of a dense-medium cyclone, a brief description may be helpful to some readers who are not familiar with its basic operation. In a typical dense-medium cyclone, illustrated in Fig.7-51 a mixture of medium and raw coal enters tangentially near the top of the cylindrical section, thus forming a strong vortical flow. The refuse moves along the wall of the cyclone and is discharged through the underflow orifice. The washed coal moves toward the longitudinal axis of the cyclone and passes through the overflow orifice, or vortex finder, to the overflow chamber. Washed coal is discharged from this chamber, usually through a tangential outlet.

Jan 1, 1991

By L. G. Austin

Introduction My first contact with industrial milling was during the time I worked in the electricity generating industry in the United Kingdom. In visits to power stations to investigate either deposits in the boiler furnaces or polluting deposits settling around the stacks. I had to check the performance of the vertical coal pulverizers, since poor pulverization aggravated both problems. Naturally, then, when I came to the USA in 1957 to take a PhD in fuel technology at Penn State, I was put to work to review the science of coal pulverization. After this reviewing, I was completely confused. On one hand, there was a well-developed understanding of stress-strain equations, and a rap- idly developing knowledge of how stressed, brittle solids fractured, based on the Griffith crack theory. On the other hand, reading in the grinding literature gave me: • Kick's Law, which was clearly not correct in the light of modern fracture theory; • Rittinger's Law, which was also clearly not correct; • Bond's Third Law of Comminution, which was claimed to have something to do with the Griffith crack theory, but where the connection between the two was made by intuitive pseudo- scientific reasoning I could not accept; • the choice of mill motor power for the most common type of coal mill, the Raymond pulverizer, was calculated from the fan power required to move air through the mill. Although I could accept the empirical connection between the two, it made no sense from the point of view of fracture energy. Even today, most books or review chapters on size reduction start from these laws. incorrect statements abound in the literature, such as “the Hardgrove Index is based on Rittinger's Law," which it is not, "The Bond theory states that work input is proportional to new crack tip length produced in particle breakage," which is not true, etc. My own test work showed that these "laws" did not fit the data for grinding of coal. At about this time, Epstein (1 948) and Broadbent and Callcott (1 956), following the original work by R.L. Brown (1941) at the British Coal Utilization Research Association, proposed describing breakage as a series of fracture stages. I took their concepts and developed the basic differential equation for a batch grinding process continuous in time, analogous to a batch chemical reactor. Robin Gardner then joined the project and did his PhD on treating batch grinding in the same way as a batch chemical reactor. He found that the basic equation had already been partially derived by Sedlatschek and Bass (1 953) in Germany. We confirmed experimentally the validity of the equations for describing batch grinding (1 962) and formulated the equation describing steady-state continuous grinding in a fully-mixed mill. At about the time this work was published, Gaudin and Meloy (1962) and Filippov (1961) independently published essentially the same equations, but without experimental proof of the validity of the concepts. I will give a brief overview of what these beginnings had led to in the design of mills for size and power, and show some of the results of this more detailed understanding of grinding processes. Concepts of Fracture Mills such as tumbling ball, rod, pebble and autogenous mills and vertical mills such as the Raymond, and E-type apply compressive stress to lumps or particles relatively slowly. Compressive stress applied to a particle of an elastic brittle solid imparts overall strain energy to the solid and produces local regions of tensile stress, (Fig. 1) (Berenbaum and Brodie, 1959). Irwin (1949) showed from solution of the stress-strain solutions that a small hole in a region of tensile stress reduces stress concentration at the hole, that is, the tensile stress at the tip of a crack or flaw in a solid is much higher than the general tensile stress in the region. The longer the crack, the higher the stress concentration. Griffith (1920) hypothesized that when the regional tensile stress is large enough, then the chemical bonds at a preexisting crack tip are stretched to breaking point, as illustrated in Fig. 2. When the bonds break, the crack becomes longer, the tensile stress concentration increases, the situation is unstable and a crack opens up (propagates) a surface of tensile stress, creating its own tensile stress at the leading edge. Stored strain energy is converted to the kinetic energy of the moving stress field, which is analogous to sound propagation through the solid, so the crack tip accelerates to velocities approaching those of sound. The moving crack will pass through regions that were previously under regional compressive stress. The equations for "ideal" and "Griffith" strengths are where a is the intermolecular distance, y is Young's modulus, g is the energy

Jan 1, 1998

By S. W. Arambur

INTRODUCTION "Mujeres Roca" is a program developed by OMA in association with the Chamber of Commerce Canada Peru inspired in “Women who Rock”, a Canadian initiative to promote a greater participation of women in the mining industry. The program includes a coaching workshop called “Kallpa Warmi”, that means “Woman Strength”, to reinforce basic skills and values and 8 mentoring sessions with top executives of the mining industry who share their experiences and knowledge with a group of 10 selected young professional women from around the country. HISTORY OMA is a nonprofit association founded in 2014 dedicated to contributing to the education of low-income university students with high academic performance in careers related to mining and earth sciences. Our mission is to empower the future professionals by strengthening their soft skills and enhancing their capacity to become leaders and agents of change in their own regions. The Chamber of Commerce Canada Peru has a long-standing commitment to gender equality and promotes women participation in male dominated industries. As it is well known, the mining sector is and has always been key to Peru’s economy, representing about 10% of the country’s GDP. Thus, mining can contribute significantly to our economic development and we believe that women can make a positive difference. A study prepared by PwC and "Women in Mining” provided indicators on the advantages that mining companies have when including women in their executive’s positions. These indicators sustain that when a mining company includes women in its management staff, its net profit increases by 6%. This information clearly shows that women participation is not only a matter of human rights and gender equality, but of productivity and profitability. Nevertheless, the mining sector has the fewer number of women on boards of any industry group in the world. Having this in mind, OMA decided to expand its activities and create together with the Chamber of Commerce Canada Peru, the program “Mujeres Roca”, in order to inspire and attract young women to the mining sector, a traditionally masculine industry. The program also aims to encourage mining companies to more actively recruit and keep women in their boards. The program was launched in 2016 within the framework of EXPOMINA PERU, the largest international mining exhibition that take place in Lima every two years.

Jan 1, 2019

By S. Hearn, S. Boskovic, A. Smethurst

"The digestion of minerals in strong acids or at elevated temperatures can result in the release of large amounts of silicic acid which can polymerize in solution to form colloidal silica and gels. Colloidal silica can cause a number of problems including interfering with flocculation, initiating crud formation in solvent extraction (SX), decreasing SX extraction kinetics and increasing SX phase disengagement times. Huntsman’s POLYSIL® polyether coagulants have a “lock and key” relationship with the surface of colloidal silica to help lower the concentration of these particles in solution, improve process efficiency and reduce reagent losses by minimizing crud formation. This paper investigates some of the issues caused by colloidal silica in hydrometallurgy processes and illustrates how the use of POLYSIL® coagulants has helped to improve process outcomes. INTRODUCTION Coagulation differs from flocculation in a number of ways. Coagulants are relatively small molecules that modify the surfaces of fine particles, weakening repulsive forces between them and allowing particles to stick together, while flocculants are generally molecules that are large enough to bridge the boundary layers between the particles. With flocculation, addition rate and mixing all make a difference to the flocs formed. Coagulation, on the other hand, is an equilibrium process. With coagulation it is only necessary to achieve good mixing, and coagulates will form as Brownian motion pushes the particles into each other. In real systems, this process is complete within seconds. Huntsman’s POLYSIL® polyether based coagulants have a strong affinity with the surface of colloidal silica formed by the polymerization of silicic acid. While silica found in minerals has a low aqueous solubility, large amounts of silica as silicic acid can be released as these minerals are digested in strong acids at elevated temperatures. The solubility of free silicic acid is about 150 ppm under pH conditions typical in acid leach environments (1), so the bulk of the silica will polymerize in solution to form colloidal silica. Colloidal silica can pass through a hydrometallurgy process without affecting any of the operations. However, it can cause a number of problems. These include gelling in a leach, interfering with flocculation or crystallization, initiating crud formation (2,3,4), decreasing extraction kinetics and increasing phase disengagement times (5,6) in solvent extraction. In these cases, coagulating the silica with POLYSIL® coagulants can reduce or eliminate the problem by enabling these particles to be flocculated and removed from the solution."

Jan 1, 2015

"There is no fast way to add more than 6.4 km (4 miles) of track, most of it underground, to the Sound Transit Link Light Rail system in the Seattle, WA area. But increasingly busy traffic through a number of neighborhoods made it necessary. That’s the challenge a joint venture of several tunneling contractors faced when starting its portion of the $1.9-billion Northgate Link Extension in 2013, and the project is still keeping it busy three years later. Completion will mean just a 14-minute ride from Northgate to downtown for an extra 60,000 passengers by 2030. The transit lines will join 57 km (36 miles) of new and under-construction light rail lines running north, east and south from Seattle.Voters tired of long commute times approved the project, a key part of Seattle’s regional mass transit system, in 2008. The extension is part of a plan adopted by the Sound Transit board in 1996 to connect the region’s major activity centers. The Northgate project is meant to relieve one of the area’s most congested traffic areas by connecting the Northgate, Roosevelt and University District neighborhoods to existing routes including downtown Seattle and SeaTac airport.Sound Transit dedicated more than a quarter of the project’s estimated expense — $440 million -— to general contractor JCM Northlink LLC to handle the tunnel segment. JCM is a joint venture formed in 2013 between Jay Dee Contractors of Livonia, MI, Frank Coluccio Construction of Seattle, WA and Michels Corp. of Brownsville, WI. The contractors regularly collaborate on large-scale projects and brought experience working on a different section of the extension they completed in 2013.JCM project leaders knew immediately that the schedule would be demanding. They had to work around the clock six days a week to meet deadlines. The extension runs through highly populated areas, meaning contractors had to follow strict city noise restrictions between 10 p.m. and 7 a.m. Plus, supervisors needed to manage the scheduling of nearly 250 employees to keep work running smoothly.“We had eight different operations and four different subcontractors trying to work in one tunnel,” said Chris Wood, JCM site project superintendent. “It’s important that we keep work linear. That’s the biggest challenge: keeping a path in and out so you can pour concrete, keep electricians and carpenters going, and accomplish everything else.”"

Dec 1, 2016

By Frank A. Aplan, Barbara J. Arnold

INTRODUCTION Coal flotation is a fine cleaning process usually present in all but the older preparation plants. It is typically used to clean either the -28 mesh (-500 pm) or - 100 mesh (- 150 pm) raw coal fines, and it is the only commonly used preparation method that can effectively recover coal particles finer than, say, -200 mesh (74 pm). To place coal flotation into its proper context, it should be noted that the bulk of the raw coal sent to preparation plants is treated by gravity concentration. On a worldwide basis, from 0 to 40% of the raw feed to a preparation plant is treated by flotation, while in the US only about 10% of the raw coal feed is floated. Roughly, 50% of US preparation plants use flotation to process at least a portion of the raw coal feed, although most of the newer plants incorporate flotation into their flowsheet. Historically, the process has only been used to treat the raw coal fines naturally occurring after mining and a minimal amount of crushing at the preparation plant. This is in sharp contrast to ore flotation where, typically, the entire tonnage is first ground and then floated. As a consequence, of the approximately 500 Mt of raw coal treated by preparation in the US yearly, it is estimated that only about 40 Mt are sent to the flotation circuits (producing -30 Mt of clean product).'O1 This small tonnage compares with about 400 Mt of ores treated annually in the US by flotation. By analogy to ore flotation, where it is an effective, large tonnage, low cost, easy to operate process, its potential in coal preparation should be great. Reasons why it has not achieved a greater potential in coal preparation have been attributed by Aplan 102 to: 1. Indifferent results in removing sulfur and ash. 2. Difficulty in floating low rank and oxidized coals. 3. Inability to make clean separations with the finer sizes. 4.Inability to clean slurries with a high percentage of clay. 5. Problems with water chemistry. 6. Competition with effective, gravity coarse-cleaning methods. 7. Low value of coal historically. 8. Dewatering problems. 9. Transportation and storage problems with coal fines. 10. Marketing problems with coal fines. 11. Inadequate attention to the process in the plant. 12. Inefficient use of capital and operating monies. Inadequate attention to procedures, machines, and processes.

Jan 1, 1991

By P. M. Jones

Introduction The success of in situ leaching depends upon: • maintaining adequate injection and recovery flow capacities in wells, and • distributing leach solution effectively to contact and leach ore minerals within the target zone. Solution balance is an important indicator of the capability of a particular in situ hydrologic design to efficiently and effectively control and distribute leach solution within the target ore zone. As part of ongoing hydrological investigations into in situ leaching, the US Bureau of Mines (USBM) is studying factors affecting solution balance in an unsaturated copper ore deposit at the Cyprus Casa Grande Mine. The purpose of this study is to assess the effect of hydrogeologic features and in situ operating conditions on changes, during leaching, in: • solution storage within the deposit, and • solution balance. At the Casa Grande Mine, In-Situ-Leaching Test Facility, solution balance estimates are obtained for 10 selected intervals during 17 months of leaching under a variety of well configurations, pressures and flow conditions. Operational and geological description The Casa Grande Mine is located 32 miles south of Casa Grande, AZ. The mining area is within the Casa Grande Deposit, a porphyry copper stock consisting primarily of an intensely-fractured, granodiorite cap overlying a copper sulfide base. As a result of mine dewatering activity, unsaturated conditions existed in the deposit prior to leaching (Schmidt et al., 1990; Friedel and Schmidt, 1991). No groundwater was present in any of the in-situ-leaching, test-facility wells after drilling and prior to leach solution injection. Also, no ground water seepage occurs in the test-facility drift. The Casa Grande, In-Situ-Leaching Test Facility consists of 58 injection and recovery wells collared in an underground drift. The wells were drilled at various angles into the granodiorite deposit in a series of 15 divergent fans (Fig. 1). The fans are 12.5 ft apart and lie parallel to one another between the 900- and 1100-level drifts. Each fan contains three to five wells drilled at angles ranging from 6° to 89° from the horizontal. The wells can operate in an injection or a recovery mode. Currently, 36 flow sensors and 28 pressure transducers (part of a hydrologic, data acquisition system) are used to continuously monitor injection-well flow rates and pressures.

Jan 1, 1993

By J. Szarkowski

"INTRODUCTION The Butte district of Montana is well-known as one of the biggest historic producers of copper in the world. It is also a classic example of a porphyry Cu-Mo system that has been overprinted by zoned, polymetallic (Cu-Zn-Pb-Ag) lode mineralization. After copper, silver was the next most important metal mined from Butte, with over 700 million ounces of total Ag production between 1880 and the present day. Although many papers have been written on the mineralogy of the early porphyry Cu-Mo mineralization and the high-grade Cu-rich lodes in the center of the district, relatively little information exists on the mineralogy of silver in Butte. The purpose of this paper is to summarize past and present silver production trends for the district, and to combine previously published work with new research to examine the mineralogy of silver in the Butte mining district. BUTTE HISTORIC SILVER PRODUCTION According to a recent USGS compilation1, Butte is one of the most Ag-rich porphyry deposits in the world, both in terms of average Ag grade (8.6 g/ton) and total production + reserves (Fig. 1). It is also the 2nd biggest Ag-producing mining district in the U.S. (after the Couer d’Alene district, Idaho). Annual silver and copper production data for the Butte district, compiled from various sources2,3,4, are summarized in Figure 2. Peak production of Ag and Cu was in 1917. While Ag production gradually declined in the ensuing decades, Cu production remained strong until the years leading up to the closure of the Berkeley Pit in 1982. In 1984, mining of the nearby Continental Pit resumed under a different operator, and continues to the present day. The Continental mine produced an average of 35,000 metric tons Cu and 575,000 oz Ag per year between 1984 and 20044. These figures do not take into account a period of mine closure in 2001-2003 due to an unexpected hike in utility prices. The long-term decline in Ag/Cu ratio of ore mined in Butte (Figure 3) is due to several factors, including: 1) changes in metal prices; 2) a downwards decrease in Ag/Cu ratio of most of the lodes with depth below surface; and 3) a shift in mining methods from selective underground mining of high-grade veins to open-pit mining of the bulk rock."

Jan 1, 2016

By Keh-Jian Shou

"This study adopts and analyzed a shield tunneling case in Taiwan, with a diameter from 5.25m to 6.70m and the smallest radius of curvature 100m, which was constructed in gravel formations as well as a vibration sensitive area. In this case study, a vibration monitoring plan, comprising the sensors inside the shield machine and on the surface, was proposed and performed to provide the vibration data during the construction. In this project, the major vibration reduction considerations were increasing the open ratio of the cutter head (to reduce the cutting function) and reducing the rotational speed of the cutter head. The results of the construction suggest the appropriateness of the above considerations. The findings could provide suggestions to the other similar shield tunneling cases. INTRODUCTION Underground vibrations during shield tunneling are typically isolated and absorbed by the soil. However, a portion of these vibrations may disseminate to the surface. These vibrations are considered micro-vibrations, which generally do not pose a threat to surface structures, and are thus seldom the focus of research. The present project was located in the Hsinchu Science Park where shield excavation bypassed numerous high technology manufacturing plants (e.g., wafer manufacturing plants). The vibrations generated by shield tunneling may affect the vibration-sensitive manufacturing equipment and instruments in these plants, and induce problems in manufacturing processes. Therefore, research to comprehensively understand the characteristics and potential influence of such vibrations is essential. To clarify the characteristics and potential influences of vibrations generated during shield tunneling, the vibration level in the excavation of the Taoyuan International Airport Access (TIAA) Metropolitan Rapid Transit (MRT) Project was examined, as it possesses similar excavation conditions to the present project. The parameters collected from relevant tests were provided as a reference to the firms along the excavation route of the present project to explain the effects of vibrations, and to relevant units for shield machine selection and project execution. No yield problems resulting from shield machine induced vibrations were reported by the firms, and the vibration levels measured in the present project were comparatively lower than those in the TIAA MRT Project. Because the related literature is scarce, this study aims to analyze the collected measurement data and provide suggestions on shield machine selection or designs. Such that we can more properly evaluate and control the impact of shield tunneling induced vibrations in similar projects."

Jan 1, 2016

By Houston Kempton, Tim Dyhr, Andy Davis, Mark Bloom, Connie Travers, Dave Atkins, Jeff Deen

INTRODUCTION Historical mining activities in the Robinson Mining District, in the Egan Range of east-central Nevada (White Pine County), resulted in five open pits (Veteran, Tripp, Liberty, Ruth, and Kimbley Pits), several large mine-waste dumps, and numerous underground workings developed for copper, zinc, gold, lead-silver, and manganese (Figure 1). Groundwater percolating downward through the copper sulfide ores produced a blanket-like supergene zone enriched in copper, which has been one of the primary objectives of mining since the early part of this century. To obtain permits for future operation, an EIS was conducted that evaluated future pit- lake chemistry, groundwater quantity and quality issues, transport of water and solutes subjacent to a proposed tailings impoundment, and the potential for waste dump acidification. AREA HYDROGEOLOGY The mine area is on a ridgeline draining to the Steptoe Valley to the east and to the White River Valley to the West. Precipitation in the Robinson Mining District ranges from a low of 15 cm (6 in.) per year in the valley to more than 50 cm (20 in.) per year in the surrounding mountains. Evapotranspiration exceeds precipitation, leading to a net loss of water from the basin as either surface flow or underflow to the north at Currie, Nevada (Eakin et al. 1967). The White River Valley, which lies to the south and west of the mine area, includes approximately 4,200 km2 (1,620 mi2) of the upper part of the Colorado River Basin. This valley is 112 km (70 mi) long, 32-48 km (20-30 mi) wide, and like Steptoe Valley, evaporation and transpiration exceed precipitation (Maxey and Eakin 1949). The local groundwater flow system in the Robinson Mining District and the proposed mine area is dominated by two main components: (1) a groundwater mound or high between the Ruth Pit and Murry Springs, and (2) a shallow groundwater flow divide that follows the ridge crest south of Ruth, which contains the Veteran and Tripp pits. Besides the alluvium, the main hydrostratigraphic units controlling groundwater flow are limestones, shales, and sandstones, along with the intrusive quartz monzonite stocks. Based on the piezometric surface measured in October 1993, three distinct groundwater regions were identified. In the Giroux Wash area, an apparent fault runs between WCC-G1 and WCC-

Jan 1, 1995

By M. A. Reyes, L. Yan

"Heat and humidity buildup within refuge alternatives (RAs) may expose occupants to physiological hazards such as heat stress. Mine Safety and Health Administration (MSHA) regulations require RAs in underground coal mines to provide a life-sustaining environment for miners trapped underground when escape is impossible. RAs are required to sustain life for 96 hours while maintaining an apparent temperature (AT) below 95°F (35°C). NIOSH tested a 10-person tent-type RA, a 23-person tent-type RA, and a 6-person metal-type RA in NIOSH underground coal mine facilities to investigate the thermal environment over a 96-hour period. The test results showed that mine air and mine strata temperatures surrounding an RA occupied by simulated miners increased over the 96-hour test period. The test results suggest that RA manufacturers should consider this increase in temperatures when calculating and evaluating RA components during surface and laboratory tests. The findings can equip stakeholders with additional considerations for calculating heat and humidity temperature profiles for RAs not tested in situ. INTRODUCTION RAs are designed to provide a life-sustaining environment for miners who cannot escape after a mine disaster. One of the major considerations for integrating RAs in mine environments is the thermal environment inside the RA and the impact that the surrounding mine conditions have on the temperature and humidity buildup. Previous research has observed that the metabolic heat of the occupants and the heat released by the CO2 scrubbing systems will cause the RA interior air temperature to increase. In addition to the increase in temperature, an occupant’s respiration and perspiration, as well as the chemical reaction of the scrubbing chemicals with CO2, will contribute to an increase in humidity within the RA. The resulting internal thermal conditions, if not controlled, can subject miners to conditions that can lead to heat exhaustion, heat stroke, or even death depending on the duration and magnitude of exposure. Apparent temperature (AT) is a temperature-humidity metric for the perceived temperature caused by the combined effects of air temperature, relative humidity (RH), and wind speed. It is used to assess the perception of indoor temperatures when workplaces are not sufficiently heated, cooled, or insulated to provide comfortable or healthy conditions. Mine Safety and Health Administration (MSHA) regulations require that RAs must be designed to ensure that the internal AT does not exceed 95°F (35°C) when the RA is fully occupied."

Jan 1, 2018

By Alexander N. Winchell

"The deposits of the east end of the Mesabi Range cover very large areas, and consist of the same taconite that produces the hematite of the central and western Mesabi. It is not deeply covered with drift, but outcrops boldly in large masses. It is not concentrated locally into enriched bodies separated by leaner areas, but is all a fairly uniform, hard, banded, unaltered chert, carrying about 25 to 30 per cent iron as magnetite.The Mesabi Iron Company, organized by Hayden, Stone & Co., and under the direction of D. C. Jackling, has taken over these deposits and is building a Concentrating plant at the new town of Babbitt, about fourteen miles east of Mesaba, Minn. This plant is to have a capacity of 2,000 to 3,000 tons per day, and is based on the application to this low grade material of the methods which have made the so-called porphyry copper properties so successful.In 1916, an experimental mill was built in Duluth; containing full sized machinery and capable of handling from 100 to 200 tons daily. A considerable tonnage of material was concentrated here during 1916, 1917 and 1918. The work showed that the resulting product could be varied, almost at will, between 60 per cent and 70 per cent iron and .006 per cent to .030 per' cent phosphorus, and that this could be made on a large scale well within the limits of cost set by iron trade conditions. A small special cargo was sent down the lakes, late in 1918, assaying 63.27 per cent iron (natural and dry) and .008 per cent phosphorus; but it is assumed that large scale shipments will normally be of a Bessemer' grade carrying about 63 per cent iron and .020 per cent to .030 per cent phosphorus. This material is a sinter, carrying only traces of sulphur and titanium, with silica between 6 per cent and 10 , per cent, and alumina, lime, magnesia, and manganese, each below 1 per cent. Some of the iron in this sinter is hematite and some is magnetite, the percentages varying according to the treatment given on the sintering machine. Like the analyses, this hematite percentage is within reasonable control. Sinter made from finely ground magnetite has much the structure of coke, being firm yet porous, and gives good account of itself in the blast furnace."

Jan 1, 1920