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By Scott Rosenthal, Dave Gilkey, Theresa Stack, Daniel Autenrieth, Ciara Kremer

Jackleg drill operators are exposed to harmful levels of hand-arm vibration (HAV). Anti-vibration handles and gloves provide modest reductions in HAV exposures and forearm muscle exertion from the use of anti-vibration handles and gloves by jackleg drill operators. The goal of this pilot study was to investigate changes in HAV with the use of anti-vibration gloves and handles compared to forearm muscle exertion experienced by operators and measured with surface electromyography (EMG). Five subjects operated the drill under four different cases: no anti-vibration controls, anti-vibration gloves only, anti-vibration handle only, and simultaneous anti-vibration handle and glove use. Muscle exertion was expressed as a percent of maximum voluntary contraction (%MVC) and was compared using Welch’s ANOVA with Games-Howell post hoc comparisons. The case with both anti-vibration controls in use simultaneously (largest grip diameter) was associated with a mean %MVC of 36.13% during operation for all forearm muscles combined, which was significantly higher than that of the other cases (p < 0.05). There were no statistically significant differences in mean HAV exposures. The anti-vibration handle with anti-vibration glove case only increased the maximum allowable exposure time by 8 min as compared to the control case without any anti-vibration controls. These results suggest that the modest HAV exposure reductions from the use of anti-vibration handles and gloves may pale in comparison to the increased muscle exertion resulting from their use, and this trade-off among jackleg drill operators is a potential concern that warrants further investigation.

Jul 12, 2021

By V. A. Evans

The preparation of a paper reporting fluorspar economics in the United States at one point in time could have been somewhat complicated be- cause of the number of producers, both large and small, that derived production from different locations throughout the Western U. S. and the Southern Illinois and Western Kentucky fluorspar districts . Companies such as Allied Chemical and William- son Mining operated in the Jamestown district of Colorado. In Walden and Salida, Colorado, the Colorado Fluorspar Company operated mines and mills.

Jan 1, 1986

By T. C. Vogt

The past history of uranium production has been dominated by strip and shaft mining methods, but recently in-situ leaching has emerged as a viable alternative for uranium recovery. (1-8) Mobil was the first company to start in-situ uranium leaching in New Mexico. Mobil&apos;s Crownpoint project areas are in Mdinley County and are part of the Grants Mineral Belt. The major uranium deposits in this area occur in the Westwater Canyon Member of the Jurassic Morrison Formation. (9-11) The ore depth at the Crownpoint project is about 2000 ft. In an earlier paper we discussed the uranium mineralogy of an areally broad sampling of Crown- point ore. (12) The organic carbonaceous material often associated with the coffinite ore was characterized in some detail. The organic material appeared to be a very mature, coaly-type kerogen derived from plant components. This result was consistent with other studies on such carbonaceous material. (13-15)

Jan 1, 1981

By Deniz Akbay

The strength of rocks and soil is a crucial design parameter in engineering projects, and it can be determined through various test methods such as uniaxial compressive strength, tensile strength, and shear strength. The point load index test is a popular indirect test method for predicting the uniaxial compressive strength of rocks. However, the reliability of the point load index’s estimation of uniaxial compressive strength of rocks is questioned due to the wide range of coefficients that are used to predict uniaxial compressive strength of rocks using point load index values. Factors such as the shape and type of rock specimen, practitioner, and test apparatus used can affect the accuracy of the point load index test. This study investigated the effect of four different specimen shapes used in point load index test (diametral, axial, block, and irregular lump tests) in predicting the uniaxial compressive strength. The rock samples were tested using four test procedures which are called diametral, axial, block, and irregular lump tests. The results showed that the irregular lump test was the most accurate in predicting uniaxial compressive strength, with the highest correlation coefficients and lowest mean absolute percentage errors. The point load index test can be used as a reliable predictor of uniaxial compressive strength of rocks when the irregular lump test is preferred.

Oct 18, 2023

By R. Mensah-Biney

The kinetics of gold-bromide loading onto activated carbon has been investigated in a batch stirred reactor. The amount of gold adsorbed by the carbon was correlated with time based on a second-order equation. The rate constant for gold-bromide species onto activated carbon ranged from 0.27 to 1.2 x 10-3 (kg/t)-l(min)-l from a solution containing 25 ppm of Au, with 0.8 g carbon per liter of solution at 30° C. The equilibrium saturation capacity (Q.) was 36±3.0 kg/t from the same system. The rate of loading increased with decreasing particle size, decreasing pH, increasing temperature, and decreasing concentration of gold in solution. The calculated activation energy E. was 8.8 kcal/mole, significantly higher than those reported for physical adsorption. The equilibrium saturation capacity (Q.) increased with increasing concentration of gold in solution, but was independent of other variables such as temperature, agitation speed, particle size and pH. These rate constants and equilibrium saturation capacity for the gold-bromide species are higher than those reported for the gold-cyanide systems.

Jan 1, 1996

By P. E. Sperry, R. E. Heuer

The United States Bureau of Reclamation is constructing the $200 million Navajo Indian Irrigation Project to irrigate 110,000 acres of land on the Navajo Indian Reservation in northwestern New Mexico. Water from the San Juan River is impounded by Navajo Dam and will be conducted for 26 miles along a series of tunnels and canals to the area to be irrigated. Tunnel No. 3 is the third in a series of six tunnels along the route. Bids for construction of Tunnel No. 3 were received by the Bureau of Reclamation in September, 1970. Fluor Utah was low bidder at $8.7 million. Ten of the 12 bids were based on excavation by tunnel boring machine. The consulting firm of A. A. Mathews, Inc., was retained by the contractor to advise on geologic and rock mechanics conditions after support problems developed in the tunnel. Tunnel No. 3 is 15,264 feet long, has an 18 foot lined diameter, rises approximately one-half inch per 100 feet in the direction being excavated, and is 6,000 feet above sea level. There is one short curve in the alignment. Minimum concrete thickness is 12 inches with eight inches from &apos;A&apos; to &apos;B&apos; line making a pay diameter of 21 feet, 4 inches. The tunnel is supported with rock bolts and steel pans. Mole excavation began on May 13, 1971 from a blasted starting chamber. After 46 feet were excavated the mole was shut down and the trailing muck handling equipment was pulled into place. After advancing 1,700 feet in 34 days, unanticipated ground water conditions and ground support problems were encountered slowing moling so that the next 1,000 feet of advance consumed 97 days until December

Jan 1, 1997

By W. K. Choi

Using electrochemical techniques it was demonstrated that bioleaching of gold from refractory pyritic ores is a complex phenomenon. Many intermediate electrochemical reactions were found to be involved in the oxidation of pyrite, using cyclic voltammetry and carbon paste-pyrite containing electrode. Furthermore, the oxidation process was controlled by solid-state diffusion as suggested by chronoamperometric and chronopotentiometric measurements. Using the Sand&apos;s approach, the activation energy of solid-state diffusion was found to be = 22.4 kJ. After bioleaching, 96% gold was extracted from the leach resi¬due with a thiourea solution in 1 hour. A generalized model for the mechanism of pyrite oxidation has been presented.

Jan 1, 1990

By A. K. Martin

Adaptive management can be described as a structured, iterative process of robust decision making in the face of uncertainty, with a goal of reducing uncertainty over time through system monitoring. Adaptive management is a tool which can be applied to not only change a system, but learn about the system (Holling, 1978). Adaptive management has been used in ecosystem management for years and guidance is provided in a variety of handbooks and publications. Adaptive management can be applied to industrial facilities and mines through the use of an adaptive environmental management plan (AEMP) and associated practices. The goal of this paper is to provide an explanation of how the plans and practices support regulatory agency approvals and permitting, span the life-cycle of mines, manage uncertainty, and assist in addressing mining impact and legacy issues. AEMPs and associated practices are useful tools to address the uncertainties in identifying, evaluating, and managing impacts of proposed new and modified mining facilities undergoing environmental review. Once a proposed project is approved, the AEMP continues as a living document, being informed, revised, and adapted by the operations. The AEMP can serve as an ongoing management tool for describing and evaluating environmental impacts, continuing in operations into post reclamation. If initiated during planning and environmental review of a project (and in some locations such as the United States [U.S.], the preparation of an Environmental Impact Statement (EIS)), the AEMP becomes a basis to identify, monitor, and manage potential project environmental impacts. Development of an AEMP for an operating mine is also beneficial since there are generally developing and ongoing environmental issues at those facilities as well. Over time, the AEMP is adapted to changing requirements and conditions including those anticipated and unanticipated. Ever increasing scrutiny of mining projects and the complexities of the EIS process highlight the uncertainties inherent in impact predictions. In making the momentous decision of initial project approval, these uncertainties can discourage the regulatory agencies? actions and create stakeholder anxiety and controversy about a project. A well-designed AEMP can support the decision to approve an EIS and enable a project to move forward. Additionally, an AEMP can inform financial assurance estimates and revisions over the life of a mine providing all parties including the citizens of the surrounding communities with greater confidence that potential environmental impacts and risks will be well managed.

Feb 23, 2014

By G. ALBORNOZ, S. Utili, A. AGOSTI, C. VALDERRAMA, R. PELL, N. Morales

The steepness of the slopes of an open pit mine has a substantial influence on the financial return of the mine. The paper proposes a novel design methodology where overall steeper pitwalls are employed without compromising the safety of the mine. In current design practice pitwall profiles are often planar in crosssection and the profile in between ramps especially tend to be planar and defined by a constant inter-ramp angle. Sometimes rock layers exhibiting different strengths require the inclination of the profile to vary with depth, but the inclination across each layer is usually constant. Here instead, a new geotechnical software, OptimalSlope, is employed to determine optimal pitwall profiles of depth varying inclination. OptimalSlope seeks the solution of a mathematical optimization problem where the overall steepness of the pitwall, from crest to toe, is maximized for an assigned lithology, geotechnical properties and Factor of Safety (FoS). Bench geometries (bench height, face inclination, minimum berm width) are imposed in the optimization as constraints which bind the maximum local inclination of the sought optimal profile together with any other constraints such as geological discontinuities that may influence slope failure. The obtained optimal profiles are always steeper than their planar counterparts (i.e. the planar profiles exhibiting the same FoS) up to 8 degrees depending on rock type and severity of constraints on local inclinations. The design of a copper mine is first carried out employing planar pitwalls, secondly adopting the optimal pitwall profiles determined by OptimalSlope. The adoption of optimal slope profiles leads to a 34% higher net present value and reductions of carbon footprint and energy consumption of 0.17 Mt CO2 eq and 82.5 million MJ respectively due to a 15% reduction of rockwaste volume.

Mar 2, 2022

By Jiang Su, Alan Bloodworth

"Composite sprayed concrete lining (SCL), consisting of two layers of permanent sprayed concrete separated by a layer of spray-applied waterproofing membrane, represents the latest development in SCL tunneling, but presents several challenges to designers. This paper aims to address the following issues: (1) how to simulate composite action, (2) behavior of a composite SCL tunnel, and (3) how to utilize composite action to achieve lining thickness efficiency. It firstly sets out the background to the composite SCL concept, and then presents a strategy for numerical investigation. The main part of the paper is the presentation and discussion of the simulation results, focusing on the lining deformation, load sharing and the safety margin of both linings. Interface stresses are also examined. It is found that significant secondary lining thickness reduction may be achieved by utilizing composite action. Challenges for the design of composite SCL tunnels are briefly discussed at the end. INTRODUCTION Sprayed concrete lined (SCL) tunnelling has seen rapid development over the last twenty years in the UK (Su 2013). One of these developments has been the replacement of traditional sheet membrane and cast in situ secondary lining with innovative spray applied waterproofing membrane and sprayed secondary lining system, which has been adopted in recent major projects, such as Crossrail (Pickett 2013), and brings many benefits to the construction (Dimmock et al. 2013). The combination of sprayed primary lining, spray applied waterproofing membrane and sprayed secondary lining is perceived to provide a structurally efficient lining system that outperforms the traditional lining system, which consists of sacrificial primary lining, sheet waterproofing membrane and cast in situ secondary lining. The claimed benefits in structural performance arises from the composite action between the sprayed primary and secondary linings via the spray applied waterproofing membrane, which produces a much stiffer cross-section and therefore a more efficient use of materials. Early investigations on the design of composite SCL tunnels were carried out without (1) definite and reliable long-term interface parameters, (2) a verified numerical modelling technique, or (3) consideration of different stress histories between primary and secondary linings (Thomas and Pickett 2011, Marcher et al. 2011). This paper reports the findings from a numerical parametric study on a typical composite SCL tunnel with varied secondary lining thickness. The results show that the true benefit of composite lining arises from the optimisation of the lining thickness (stiffness) ratio between the primary and secondary linings to achieve secondary lining safety and thickness efficiency at the same time."

Jan 1, 2016

By E. D. Attanasi, M. W. Bultman, J. H. DeYoung

The 1987 recovery in the US mining industry permitted modest increases in exploration expenditures that year (Fig. 1). Favorable market conditions apparently allowed the gains to be sus¬tained in 1988. There was great uncer¬tainty early in the year as to whether world and US economic growth would be a casualty of the October 1987 worldwide stock market crash. By nearly all measures, the performance of the US minerals industry was strong¬. A weak US dollar and the strong growth of worldwide capital spending resulted in in¬creased US exports of con¬sumer and producer durable items as well as strong domes¬tic demand for metals. In fact, 1988 US steel industry exports reached levels not attained since the mid-1970s. It was re¬ported that steel industry labor productivity is now roughly twice the 1981 level. Among metals with signifi¬cant average nominal price increases in 1988 over 1987 levels were copper (43%), alu¬minum (52%0), and zinc (44%). Even though domestic copper production has grown since 1984 by 30% in mining and 24% in refining, em¬ployment in mining and milling is 4% below the 1984 level. Domestic unit costs are reported in the range of $1.10 to $1.33/kg (50 to 60 cents per lb). The US base metals industry is proba¬bly in its best cash position in many years. However, there has not been a large increase in exploration expendi¬tures. This is due mostly to the percep¬tion that worldwide capacity can easily be increased. Industry reports do show that "low cost" copper producers have been buying reserves from weaker com¬panies. The 1987 Society of Economic Ge¬ologists (SEG) survey of exploration activity indicated that 95 cents out of every exploration dollar spent by US and Canadian firms was for base and precious metals (Economic Geology, January 1989). Press reports, though, make it clear that most of that spending was targeted for precious metals. US claim staking data (Fig. 2) rein¬forces this suspicion. The apparent lag in new claim filings in Alaska and phe¬nomenal growth in filings in the lower 48 states suggests that exploration is directed at relatively low risk targets. In the conterminous 48, more than 40% of the new claims were made in Nevada, indicating an emphasis on gold (although even Nevada had some base metals exploration). Even with a price decline during 1988, the difference between the gold price and average global "cash" production cost of $7.30/g ($227 per oz) still leaves room for an attractive return (Northern Miner, 1988). Exploration plans and results Metals Economics Group (Halifax) indicated in its 1988 exploration survey that exploration expenditures by US and South African firms would be at about 1987 levels. It also reported that Canadian and Australian firms were ex¬pected to reduce their exploration spend¬ing significantly. Firms in both coun¬tries were caught in changes in tax laws. In Canada, the Mining Ex¬ploration Depletion Allow¬ance (MEDA) once permitted generous tax write-offs to in¬dividuals investing in miner¬als exploration. It was to be replaced in January 1989 with the Canadian Exploration In¬centives Program (CEIP). Although structured and ad¬ministered differently than MEDA, CEIP can provide an equivalent or possibly greater financial incentive for min¬eral exploration, for junior mining companies. In 1987, it is estimated that junior com¬panies accounted for two¬-thirds of the off-property (or grass roots) exploration ex¬penditures in Canada. Australia will revoke the tax exemption for gold in 1990. Australian operators, therfore, were expected to concentrate efforts on speeding up pro¬duction during 1988 and 1989. In 1988, Chile and Mali announced special efforts to attract gold explora¬tion by foreign companies. The Cana¬dian government continues to express concern about the write-off (and drawdown) of base metal reserves. Falconbridge did announce a new initiative to identify additional base metal reserves in the Sudbury district. Discoveries Gold deposits dominated discoveries

Jan 1, 1989

By Gordon Miner

GENERAL DESCRIPTION At the Star mine, Hecla Mining Co., located in Burke, Shoshone County, ID, the mining method is horizontal slicing (timbered or untimbered) cut-and-fill using hydraulic sand backfill. Back stoping as well as breast stoping is used; see Fig. 1. The annual produc¬tion from all cut-and-fill mining is 259 816 t (286,400 st). The average depth ranges from 366 to 2408 m (1200 to 7900 ft). A majority of the tonnage comes from an average depth of 2240 m (7350 ft). Average travel time to the stope is 1 hr. GENERAL ORE BODY REQUIREMENTS AND LIMITATIONS Size and Ground Conditions The ore body is 610 to 1219 m (2000 to 4000 ft) along the strike and 2408 m (7900 ft) on the dip, with an average vein width of 2.6 m (8.5 ft). The dip is essentially vertical. The ore body is heavily sheared and broken on the main fault structures; it is more competent on the wall veins but everywhere subject to rock bursting at depth. The hanging and footwalls are strongly fractured on the main fault structures, again more competent but still fractured on the wall veins. All host rock is Revett quartzite. Value and Production Requirements The ore value is $56 per ton. Production require¬ments are 1089 t/ d (1200 stpd), five days per week. Why Cut-and-Fill? The slusher extraction method is the most satisfac¬tory and practical for narrow vertical veins with sheared walls. The method must be quite selective to be com¬mercial. The veins are too narrow for rubber-tired ex¬traction and the grade is not high enough to make resuing practical. DETAILED DESCRIPTION Development Access to the main shaft is by the 3.2-km (2-mile) main haulage tunnel. The level intervals in the vertical timbered shafts are 61 m (200 ft). Vein development is by crosscut and laterals located 30.5 m (100 ft) into the vein wall. Vein crosscuts are on ±61 m (200 ft) centers. Access to the vein is by raises carried up with the stope. The raises are equipped with four wood¬cribbed chutes (two orepasses reinforced with angle irons, one manway, and one timber slide, Figs. 2-6). The initial cut is located 6.7 m (22 ft) above the main level (Fig. 1). Stoping Cycle Detail Sampling: Grab samples are taken from the mine cars. Channel sampling of the faces and backs is done when necessary. Drilling: For stope drilling, Ingersoll-Rand 300M percussion drills are used, with Ingersoll-Rand 300M and Gardner-Denver RB83M stopers, used for back stoping. Breasting down is done with Gardner-Denver S83FM jacklegs. The back stoping rounds are 0.9 to 1.8 m (3 to 6 ft) and the breast stoping rounds 1.5 to 3m (5 to 10 ft). Blasting: For stope blasting DuPont Tovex 100 watergel with fuses and igniter cord is used primarily. Approximately 60% of the blasting is done with am¬monium nitrate-fuel oil (ANFO). For development headings and shaft sinking, DuPont Drivex watergel with Ensign Bickford Nonel Primadet primers is is used. Mucking: Broken ore is removed entirely by slusher in the stopes. These are mostly 3.7-7.5-kW (5-10-hp) Ingersoll-Rand air-powered or 7.5-37.3-kW (10-50-hp) Ingersoll-Rand electric-powered. In the horizontal de¬velopment headings Eimco 12B and 21 and Atlas Copco LM-56 overshot mucking machines and a 0.76 (1 cu yd) hydroelectric Wagner HST-1 LHD (load-haul-dump) unit are used. Rockway for Extraction: The cribbed chutes (Fig. 7) are nearly vertical and hexagonal in shape and are made of spaced 152 x 203-mm (6 x 8-in.) timber, reinforced with varying weights and sizes of angle iron, depending on wear requirements. Rectangular cribbed chutes are used where ground pressure is not a problem. Cribbing is added in 2.4 or 3-m (8 or 10-ft) extensions accord¬ing to the floor height (whether timbered or untimbered). See Figs. 8 and 9. Backfilling: Backfilling is done with hydraulic sand consisting of classified mill tailings at 65% solids. The fill contains approximately 20% 325-mesh fines. The tailings are pumped 3353 m (11,000 ft) at 50-55% solids 22.7 dry t/h (25 dry stph) with a duplex piston pump. Cycloning increases the tailings density to 65% solids. Fill storage is in a 998-dry-t (1100 dry st) agi

Jan 1, 1982

By P. Moran

A promising and increasingly popular tailings management strategy is to place cemented tailings underground as backfill. The placement of cemented tailings results in enhanced geotechnical stability and a purportedly more geochemically stable tailings form with lower leaching potential. Conventional subaerial tailings management can negatively affect water quality if weathering causes acid rock drainage and metal leaching (ARD/ML) to occur, whereas cemented tailings can limit leaching via porosity and permeability reductions and contribute alkalinity to limit the potential for ARD/ML. Cemented tailings backfill appears to be better, but how can this advantage be proven? Currently, cemented tailings backfill is utilized at mining operations internationally, but there is limited guidance on performing geochemical characterization and associated evaluation of the potential impacts to water quality. As a result, various approaches to simulating the leaching behavior of cemented backfill have been utilized, including some methods that are unlikely to represent weathering under site-relevant conditions. American Society for Testing and Materials (ASTM) Standard Test Method C1308-08 is a diffusion testing method initially developed to assess constituent release rates from solid nuclear waste forms. ASTM C1308-08 uses intact cylinders that more closely represent actual backfill than methods that require crushing or size reduction. It is widely believed that leaching from actual cemented tailings is controlled by diffusion through small pores that exhibit limited exposed surface area, making diffusion testing particularly applicable to cemented tailings placement.

Feb 27, 2013

By Shuchen Sun, Ganfeng Tu, Chuan Liu, Xiaoping Zhu

The hydrometallurgical residue generated during the hydrometallurgical processing of waste automotive catalysts is valuable as a secondary resource. In this study, a process of platinum recovery from the hydrometallurgical residue by high-temperature smelting process was proposed. Results of lab- and pilot-scale experiments show a high recovery of more than 98% of platinum and glass slag was successfully obtained by controlling the Fe3O4 addition, C/Fe3O4 mass ratio, CaO/SiO2 mass ratio, smelting temperature, and holding time at 15 wt%, 0.4, 0.6, 1500 °C, and 120 or 60 min, respectively. Results from the pilot-scale show that the platinum contents in the glass slag and alloys were less than 4 g/t and more than 0.6 wt%, respectively. Therefore, the process proposed in this study is a prospective and effective approach for reutilization of hydrometallurgical residue containing platinum.

Oct 12, 2020

By S. K. Sharma, ANKUSH GALAV, G. S. P. Singh

Providing a suitable protective water barrier pillar (PWBP) is common to reduce inundation hazards in underground coal mines. The imperative factors influencing its performance include the water head acting on the pillar, cover depth, pillar width, strength properties, and permeability characteristics. The mechanical failure of such pillar is a stress-controlled phenomenon, whereas the hydraulic failure is a strain-based phenomenon. A finite-difference numerical modeling approach was developed to study the hydro-mechanical coupled behavior of protective water barrier pillars. The mechanical stability was evaluated in terms of the percentage of failed (ZoF) and intact zones. The influence of the strain-controlled weakening on the permeability of the flow medium was studied through the coupling of the mechanical and hydraulic effects. The coupled steady-state model was used to estimate the outflow rate and its hydraulic stability. The adequacy of the protective pillar was also investigated by assessing mechanical stability and capability to resist hydraulic pressure against the maximum expected water head. A seepage rate-based classification system has also been proposed to evaluate the seepage potential and assess the hydraulic stability. The model has been validated for two case studies at the cover depth of 136–189.5 m and the existing pillar width of 16–42 m against the water head of 25–141 m.

Aug 18, 2022

By John Hillman

The reconstruction of the British mining industry which began shortly after nationalisation in 1947 continued until well into the 1950s with the support of the Government. This support diminished as plentiful supplies of cheap oil from the middle East made many coalmines un¬economic. The planned rundown which followed was foreshadowed in the 1965 White Paper which outlined the financial measures to speed the closure of uneconomic collieries. The decline appeared likely to continue into the 21st century as the role of coal in UK energy requirements became less and less significant. This policy was well on course in 1973 when the oil crisis led to a complete reappraisal of the position. A Government examination in 1974 of the coal industry then supported the National Coal Board&apos;s general strategy of major capital investment as set out in its &apos;Plan for Coal&apos; (1). Plan for Coal aimed at halting the intended decline in the industry&apos;s capacity and restoring it to 120 million tons of deep mined output in 1985. The Plan proposed action along the following lines (a) increase output by major Investment in construction at existing collieries with sufficient reserves to give 13 million tons per annum extra output (b) invest in extending the life of existing collieries by gaining access to additional reserves, to give 12 million tons per annum extra output © by sinking new mines, produce an extra 20 million tons per annum. Figure 1 shows the contribution (2) expected from each of the three sources up to 1990/1. The measures taken were expected to offset the inevitable exhaustion of existing pits and provide the additional 40-42 million tons capacity needed to restore annual output in 1985 to 120 million tons. Parallel to this programme for expansion of deep mined output, opencast mining was to produce 10-15 million tons per annum. A recent review however, indicates that new mines are likely to yield only 10 million tons per annum by 1985, mainly owing to the need for public inquiries into proposals for new mining and the resultant fact that planning approval takes &apos;much longer than originally expected. Nevertheless the objective of 120 million tons deep mined output by 1985 is still considered achievable of means of a greater contribution from existing mines. This programme was originally estimated to cost 2,-56 million dollars at March 1974 prices, out revised estimates of the cost of Plan for Coal indicate that it may now exceed 6,960 million dollars at mid-1977 prices. Plan for Coal includes a substantial investment in new and reconstructed coal preparation plaints. At mid-1977 prices more than 322 million dollars Ls so be invested in coal preparation projects in the UK, at the rate of, about 70 million dollars a year. This rate 3f capital expenditure has been maintained over the past three years and covers investment in a wide range of coal preparation coals for the power station, coking and domestic markets.

Jan 1, 1978

By V. T. McLemore

NMBGMR has been collecting data on uranium mining districts, mines and mills since it was created in 1927 and has converted years of historical data into a relational database that can be imported in ArcMap. The database includes information on mining districts, mines, mills, geochemistry, photographs, and bibliography. The available data includes location, production, reserves, geologic, geochemical, resource potential, and other data. The NURE stream-sediment data provides geochemical analyses for >27,000 samples collected in New Mexico during 1970s. The NURE hydrogeochemical data provides analyses for water samples. ArcMap includes location of individual ore bodies as polygons and incorporates the mines and NURE data as individual site locations. The purposes of these databases are to provide computerized data that will aid in identifying and evaluating resource potential, resource development and management, production, and possible environmental concerns, such as physical hazards (i.e. hazardous mine openings), indoor radon, regional exposure to radiation from the mines, background geochemical data and point-sources of possible pollution in areas of known mineral deposits.

Jan 1, 2010

By Aidan Giblett

"It is important to make a clear distinction on the scope of work of the Global Mining Standards and Guidelines (GMSG) guideline for sampling grinding circuits. Many mineral processing practitioners are familiar with the sampling and surveying of comminution circuits, in particular grinding circuits, for performance diagnosis and the identification of optimization potential. These surveys often provide information for population balance modelling of individual comminution and classification unit processes, and consequently require a large amount of information to calibrate and validate the unit models. As a result, these surveys are labour intensive, somewhat expensive, depending on internal capabilities of the mine site or company and have a long time span from initial planning to final reporting. They also require a high level of expertise to process the survey data, simulate the circuit and make sensible recommendations on follow up actions.As a result of these factors, full grinding circuit surveys are not a common occurrence at a large number of processing facilities, and several years may pass before follow up surveys are conducted. In the ensuing gap between surveys, the efficiency of the grinding circuit must often be taken on faith, and this is considered an unacceptable lapse in an environment of strained natural resources and low commodity prices.To support the rapid and productive energy efficiency analysis being promoted by the ICE working group, a much simplified survey and analysis process is required. The process reflected in the GMSG guideline is indeed more conducive to frequent analysis, at a skill level that is well within the capabilities of the average plant metallurgist. As a result, it becomes feasible to conduct mini-surveys of plant performance, which can be used to quickly and precisely diagnose the high level impact of process changes. These methods can also be manipulated by an experienced practitioner to provide sufficient information for modelling of the circuit to estimate the impact of process changes, allowing an initial rate and rank exercise to focus plant optimization efforts."

Jan 1, 2015

By T. Karalis

Perlite is a generic term for naturally occurring siliceous rock. It is a form of natural volcanic glass the main constituent of which is amorphous silica (70-76%). Perlite, when heated rapidly to its softening temperature range, expands from four to twenty times its original volume, due to the presence of 2-6% ?combined? water, in a manner similar to popcorn as the H2O molecules vaporize and create countless tiny bubbles, which account for its exceptional physical properties: light weight (30-150 kg/m3), low thermal conductivity, cellular structure, in general. These properties make perlite adaptable for numerous end-uses, which can be distinguished into 2 categories: Expanded perlite applications in: a) Construction Materials market (e.g. Acoustical Tiles, light-weight insulating Mortars, self-levelling Concrete), b) Agriculture (e.g. Hydroponics), c) Food and Chemical industry (e.g. Filtration Media, Filler, etc.), d) Insulation and Heat resistance technology (e.g. Cryogenic vessels, Refractory Bricks, etc.). Unexpanded Graded perlite applications in: a) Phosphoric Fertilizers industry (additive in H3PO4 production), b) Foundry technology (slag-aid, insulating and heat resistance agent).

Jan 1, 2006

By Howard L. Hartman

The federal government&apos;s declining role in funding research has created a critical dollar gap for university departments of mining engineering. Industry is not unaffected by the drought: not only is research lagging, but hard- pressed universities are going to the private sector for funding. This paper&apos;s thesis is not that mining companies should shower universities with philanthropy, but that a quid pro quo ought to be exacted: a research partnership should be contracted on a pay-as-you-go basis. A model exists at The University of Alabama (albeit not in mining) that may lend itself to adoption by our industry.

Jan 1, 1984