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By P. -É. Boileau, S. Mandapuram, Shiping Ma, S. Rakheja

The absorbed power (PAbs) has been suggested as a better measure of human responses to whole-body vibration, since it relates to the cumulative energy dissipated by the body exposed over a given duration. Moreover, unlike the other measures, the PAbs can adequately account for the intensity of exposure. Although, the vast majority of off-road vehicles impose considerably severe vibration along the horizontal axes, the vast majority of studies on biodynamic response characterization consider only vertical vibration. Only a few studies have reported PAbs responses of the seated human body exposed to horizontal vibration and the major contributing factors [1]. This study aims to characterize the PAbs responses of seated human subjects to horizontal (uncoupled x- and y-axis) vibration as functions of the vibration intensity, subject mass, seat height and the, type of back support. Methods Experiments were conducted using a rigid seat with an adjustable backrest inclination and seat height. The seat was installed on a horizontal vibration simulator and the forces at the seat base and the backrest were measured by three-axis force plates. Two single-axis accelerometers were installed on the seat back and the platform, oriented along the axis of motion. The experiments were performed using three different seat heights (350, 390 and 410 mm), back support conditions (NB- no back support and sitting erect; Wb0- Upper body supported against a vertical back support; and WbA- back supported against an inclined backrest, while sitting relaxed) and three different magnitudes of broad band excitations in the 0.5-10 Hz frequency range (0.25, 0.5 and 1 m/s2 rms acceleration under x-axis and y-axis, applied independently). A total of 8 healthy adult male volunteers with total body mass ranging from 59.4 kg to 92 kg and aged between 21-51 years took part in the experiments. The subjects were seated with their hands in lap, and feet supported on the moving platform for each posture. Each measurement was repeated 2 times, while the data were analysed using a bandwidth of 50 Hz and frequency resolution of 0.0625 Hz. [ ] The data were analyzed to derive the absorbed power characteristics of the body at both seat pan and backrest interfaces, while the coherence among the measured forces and accelerations wer

Jan 6, 2006

By Felix Torres

CVRD is one of the largest mining companies in the world. Adoption of new technology and improving operational efficiency are top company priorities. Initiatives in these areas have driven continuous company growth and an increased share price. During the first half of 2001, the company had a net profit of nearly US$ 500 M.

Jan 1, 2003

By J. C. Johnson

High-grade ore found in a shaft pillar at the Homestake Mine prompted a request for a Bureau of Mines study to determine if selected areas of the shaft pillar could be mined without jeopardizing the shaft. The design approach was to perform a finite-element study of the shaft pillar. Plan view and vertical sections were developed using two-dimensional, elastic, plane strain assumptions. Input into the model consisted of in situ stress measurements geologic mapping rock mass properties, and previous mining history. Results from the study indicated that ore-beating sections of the shaft pillar could be removed within the displacement tolerances acceptable for the shaft. Homestake engineers, using the results of the finite-element study and considering other factors, have designed and are beginning to mine the shaft pillar.

Jan 1, 1989

By Richard Mainiero, Donald A. Hurd, James H. Rowland

In previous conferences the authors have reported on toxic fumes generated by the detonation of ANFO. The research reported here extends the earlier work to include an emulsion blasting agent and ANFO/emulsion blends. Explosive mixtures were shot in 4-inch Schedule 80 steel pipe in a chamber in the experimental mine at the Pittsburgh Research Laboratory (PRL). Following each shot, the fumes in the chamber were analyzed using on-line instrumentation. A major goal of the research was to gain a better understanding of the factors that lead to the generation of toxic fumes in blasting operations. Earlier studies have suggested that the high nitrogen dioxide and nitric oxide concentrations in product clouds might be the result of the poor confinement provided by relatively weak ground strata or the exposure of the explosive to ground water prior to shooting the shot. Various mixtures of ANFO and emulsion were detonated in schedule 80 steel pipe and galvanized sheet metal pipe to evaluate the effect of confinement. Explosive mixtures were also allowed to soak in water for less than one day, one week, one month, and two months to determine which explosive mixtures would be degraded and observe what effect this degradation had on fume production. Results indicated that the production of nitrogen dioxide increased with low confinement of the detonating explosive and with exposure of the explosive to water.

By Mann V. I

The Mount Isa property has been in operation since June, 1931; during this period operations have changed from the production and treatment of carbonate ores to the mining, milling, and smelting of sulphide ores. This change involved revision of ore-extraction methods.Stoping methods have been designed to apply to parallel lodes ranging from 10 ft. to over 100 ft. in width and dipping at 50° to 60°. The hardness and strength of the ore and enclosing wall rocks, together with the strength and thickness of the country rock between lodes, have permitted the selection of a sub-level stoping method for mining the sulphide ores.

Jan 1, 1938

By F. Mielli

Mine operation managers are constantly tasked with identifying new ways to produce more, while using less. Mine effectiveness and reliability is the foundation of improving operational performance. Timely decisions need to be made in order to identify production waste. Intelligent data needs to be visualized in real-time. Operational events need to be captured with impacts quantified and root causes identified. Unfortunately, complex information technology solutions often have a very poor success rate, with some reports suggesting 30 percent of complex projects are canceled before completion.

Feb 27, 2013

By Shosei Serata

The application of conventional rock mechanics methods to complex ground is very difficult. Yet. In most underground mining, the ground is complex. Therefore. conventional methods are not only difficult to apply but are also generally ineffective in improving underground operations. Similarly, conventional rock mechanics instrumentation has not significantly contributed to improvement of mining practice. Rock mechanics instrumentation in underground mining operations is often considered to be" merely an expensive nuisance. A new system of mine instrumentation is changing this traditional view. The new instrumentation method, called the SPDR method, was originally devised by Dr. Shosei Serata to meet the specific requirement of establishing the Stress Control Method of mining upon a solid quantitative foundation. This instrumentation method has contributed substantially to the widespread success of Stress Control mining under widely varying ground conditions in different types of underground mines (1.2.3.4). The instrumentation method provides an effective means for optimizing Stress Control mining patterns, simultaneously achieving reduction in production costs and substantial improvements in safety (5). The principles and methods of the new instrumentation method are described below, with field examples.

Jan 1, 1986

By C Ferron, R Massimi

There is approximately 240 000 tonnes of arsenic trioxide dust stored in underground chambers at the Giant Mine in Yellowknife, Northwest Territories, Canada. To minimise the amount of arsenic released to the environment, a number of closure alternatives are being considered. One alternative is to mine the dust, encapsulate it and dispose the mixture into a secured landfill. The potential leaching behaviour of the encapsulated dust was examined in a laboratory programme. Cylinders of arsenic trioxide dust mixed with cement/aggregate and bitumen were prepared. The dust content varied between zero per cent and 60 per cent of dry weight charge. The cement cylinders were tested for strength, leaching and freeze/thaw durability. The bitumen cylinders were tested for leaching only. The results show that increasing dust content in the cement cylinders negatively impacts the strength and the freeze/thaw durability of the samples. A contributing factor is likely the high water to cement ratio in the samples. Leaching tests show that arsenic release from cement cylinders increase with increasing dust content. Arsenic concentrations in leachates do reach a plateau at a certain point in time. Arsenic release from bitumen cylinders was significantly lower compared to cement cylinders. Arsenic concentrations eventually were less than 0.05 mg/L in samples containing up to 25 per cent arsenic trioxide dust. Arsenic leaching rates were found to decrease by an order of magnitude over the course of three months.

Jan 1, 2003

The Study of Jamesonite Flotation Behaviour

Sep 13, 2010

By Simon D. Strauss

This article is part of a presentation Strauss made Dec. 3, 1988 at the sixth annual American Mining Hall of Fame (AMHofF) banquet in Tucson, AZ. At the banquet, the Mining Club of the Southwest honored Strauss by naming him the 1988 recipient to the AMHofF. (Industry newswatch, page 6). Introduction In the last two decades, there have been two previous periods of sharply rising copper prices. One was in 1973-74; the other in 1979-80. In both cases, a short period of high copper prices was followed by steep and prolonged periods of low prices. In the language of the market analyst, the high prices were called spikes. Is it likely, then, that three or four years from now the present price of copper will show up as a spike in the charts? There are substantial differences between the current position and the markets of 1973-74 and 1979-80. In 1973-74, US copper prices were subject to government-imposed price ceilings of $0.60 per lb through early December, 1973. These were increased to $0.68 per lb through May, 1974. Meanwhile, mines outside the United States were free of restraint. Rising demand in London pushed the price above $1 per lb in the third quarter of 1973. In early April 1974, it reached a peak of $1.50 a lb. In May of that year, the US ceilings were removed and the domestic price began to rise. What were the causes of this price strength? While ceilings held the domestic price below the world market, domestic consumers acquired all the tonnage offered by US mines. Meanwhile, in the world market, buyers feared prolonged shortages. The Club of Rome had warned that world economic growth would be limited by impending depletion of mineral resources. There was al so concern about a copper cartel that might increase copper prices even more. In early 1974, at a special United Nations session, developing-country spokesmen warned the industrialized world to expect high commodity prices. Also adding to the concerns of copper buyers was the fear that when US wage contracts expired in mid-1974, protracted strikes might ensue. The combined effect of this atmosphere was that consumers accumulated large amounts of copper, far above their normal carry. By midsummer, the position changed radically. While some strikes did occur in the United States, they proved short-lived. The most significant development was a massive economic recession. The energy crisis was causing enormous dislocations in the world economy. Construction activities were sharply reduced and automobile sales plummeted. Public confidence was shaken. Copper sales fell precipitately as consumers decided to draw down their swollen inventories. The price came down too. In April 1974, the London price- as noted-was $1.50 per lb. By the end of the year, it was down to $0.57 per lb. In 1973-74, the copper price rise was fed by fears of shortages, cartels, and strikes. In 1979-80, the market rose on the strength of good demand plus political and military instability. The price rise was also fed by inflationary trends, exemplified by speculation in precious metals. Prior to this second spike in copper prices, the market had been depressed despite economic recovery. Despite reduced US output, copper production in the rest of the world was increasing. Though consumption was rising, surplus stocks reached unprecedented levels. At the end of 1977, the aggregate of stocks at the refineries and in the exchange warehouses was 1.5 Mt (1.67 million st). That was more than three months' shipments at then rates of consumption. In 1978, the statistical position improved. Shipments of refined copper that year were 8% greater than in 1977. The result was reduced visible stocks but -just as was to be the case in 1984-86 - the price remained depressed. In 1979, an even sharper drop in visible stocks occurred as shipments rose by another 4%. This time, though, the price responded. In December 1978, the price in current dollars had been less than $0.70 per lb. By the end of 1979, it was above $1 per lb. And in the first quarter of 1980, it averaged more than $1.20 per lb. Military and political developments were a factor. The hostage crisis in Iran was followed by the Soviet invasion of Afghanistan. To many, it appeared that war might result if the Soviets pushed on into Pakistan. At the same time, speculation drove gold and silver prices to incredible levels. As silver prices rose above $40 per oz), copper above $1 per lb did not seem unreasonable. Then, in March 1980, the precious metals price boom collapsed. Copper prices also retreated, though not as much. In February 1980, the copperprice averaged $1.32 per lb. By December, it had dropped to $0.85 per lb. In 1981 and 1982, the decline continued. The present copper situation Why is the current situation different from these two previous episodes of high prices? First of all, today's inventories are far lower than they were in either 1974 or 1980, when the price declines began. The stocks at refineries and in exchange warehouses are substantially less than they were then. More significantly, consumers today have very limited holdings compared with stocks they carried in 1974 and 1980. Even if faced with a downturn in their own business, copper consumers will not be able to stay out of the market. To the extent they have any business at all, they will have to buy copper. And on the supply side, there is also a significant difference. The year 1988 has been unusual. US production is up sharply, responding to higher prices. But production is little changed in the rest of the Free World. Indeed, several of the major producing countries have mined less copper this year than in 1987 despite the prices increases. Peru has had labor problems that have cut its output. Zambia and Zaire suffer from shortages of expatriate technicians, not to mention spare parts. They also suffer from the failure to maintain equipment or perform adequate development. In Chile, the two largest mines, Chuquicamata and El Teniente, have not met their production targets. Chuquicamata is struggling with high arsenic content in its ore. And, in late 1987, severe rockbursts hampered operations at El Teniente. In addition, depletion of some mines in Australia and South Africa has offset rising output at other properties. Indeed, when final mine output figures for 1988 become available, it may well be

Jan 1, 1989

By Robert M. Miller, Judy Milligan

A computerized "Decision Support System" was used to evaluate an "Underground Platform Approach to Oil Recovery". Over a billion calculations were required to model the mine layouts; learning curves for shaft, tunnel, and well drilling scenarios; detailed cost, schedule, and resource data; complex oil production and material balance data; varying reservoir parameters; and complex oil pricing curves. Simplified "optimum" solutions were defined in terms of muck hoisting capacity, barrels of oil per day, and present worth. Complex "optimum" solutions were defined in terms of minimal exposure scenarios during development followed by shortest time to full production.

Jan 1, 1987

By R. H. Colquhoun, A. W. Bolland, H. J. H. Pieterse

"American Barrick Resources Corporation acquired the goldstrike property in north eastern Nevada in early 1987. Since that time, the company has undertaken an extensive development program. This paper will overview the milling, autoclaving, and heap leaching facilities which are part of the $400-million Betze development plan. Gold production at the Mine has increased from 40,000 ounces per annum in 1987 to more than 500,000 ounces in 1991. And in 1992 Goldstrike will be the single largest North American gold mine when it produces 1,000,000 ounces.1. SUMMARYAmerican Barrick Resources Corporation acquired the Goldstrike property in northeastern Nevada in early 1987. Since that time, the company has undertaken an extensive development program. This paper will overview the milling, autoclaving, and heap leaching facilities which are part of the four year, $440- million Betze Development Plan. Gold production at the Goldstrike Mine has increased from 40,000 ounces per annum in 1987 to more than 500,000 ounces in 1991. In 1992, Goldstrike will be one of the largest North American gold mines, when it produces appro approximately 1,000,000 ounces.2. INTRODUCTIONAmerican Barrick Resources Corporation is in the process of developing the Betze gold deposit, located on Barrick's Goldstrike property in the Carlin area of Nevada, U.S.A.The Goldstrike Mine has 17.9 million ounces of proven and probable reserves and an additional 460,000 ounces of possible reserves. Additionally, the Purple Vein deposit, one mile north of the Mine, is expected to add at least 4 million ounces of reserves by year end 1991.The Goldstrike ore body contains both oxide and sulphide ore. The largest part of the reserves are contained in the high grade, deep sulphide ore. Mining of the sulphide ore contained in the Betze deposit involves the removal of large quantities of overlying oxide ore. Oxide ore is treated through one of the following processes:-"

Jan 1, 1992

A mathematical model of a decision is represented by the table in Figure 4. The numbers in the matrix are a measure of the effectiveness of each alternate for an objective. In this case, it is the per cent of desired profit that is satisfied by each alternate for each objective. By adding all of the values in a row, we have a measure of which alternative is best for both objectives. However, if you think that alternate 1 is necessarily better than alternate 2, you are wrong. Why? Because the weight of the objectives may not be equal. It could just be that this plant can sell more than it can produce, so that the loss in profit represented by increased sales is more than twice as important as reductions in costs. (See Figure 5). In that case, let us weigh objective 1 as 30 per cent and objective 2 as 70 per cent. It must also be obvious that the sum of all the "objective weights" must equal 100 per cent. Now, by evaluating our alternates according to their weights. we find that alternate 2 is better than alternate

Jan 1, 1964

This three-part report provides the results of an industrial engineering study that assessed the safety and efficiency of underground ore handling in U.S. metal mines. One of the primary findings of the study was the improvement needed in controlling muck size in the are handling stream. Oversize ore is causing much additional labor, creating work stoppages, and is a principal cause of accidents in metal mines. The specific objectives of the Present Systems Study were to determine and evaluate the productivity of existing ore handling systems in a representative group of underground metal mines, to relate the available historical accident record statistics of these selected mines to the specific operations in the overall ore handling operations, and to highlight the potential areas of improvement for each mine on the basis of productivity and associated hazardous index for each operation.

Jan 1, 1974

By D. H. Gurinsky, D. G. Schweitzer

The empirical equilibrium constantsd the heat of reaction for the reduction have been determined from 300° to 500°C. The mechanisms of the oxidation of uranium and magnesium from dilute bismuth solutions have been investigated. The kinetics of some of the reactions were affected by air adsorbed on the uranium oxides. THE Liquid Metal Fuel Reactor studied at Brook-haven National Laboratory uses a solution of uranium in bismuth as the fuel. Corrosion and stability problems require that the fuel solution contain corrosion inhibitors and deoxidants. Of the additives tested, magnesium has been found to be the best deoxidant and zirconium the best corrosion inhibitor in bismuth. Some experiments1 indicate that magnesium may play a secondary role in increasing the effectiveness of zirconium in inhibiting corrosion. The work reported here was intended to determine the relative concentrations of Mg and U required to prevent oxidation of the uranium fuel in the event of an air leak. In addition, an attempt was made to identify some of the mechanisms of the oxidation and reduction reactions occurring in the liquid bismuth solutions. EXPERIMENTAL Apparatus and Procedure—The equipment shown in Fig. 1 was designed so that the reaction kinetics could be studied as functions of temperature, melt composition, pressure, and flow rate. At the start of a run, the 10-liter bell jar (F) is removed by opening the bottom Dresser fitting (L), and the bismuth and additives are placed in a pyrex crystallizing dish (I). The bell jar is replaced, the system is evacuated to 10-5 to 10-6 mm Hg and then the heater (J) is turned on. When the selected temperature is reached, the oxidizing gas is admitted to the system to the desired pressure as read on the manometer (D). Gas is introduced through the slide tube (Q) which has a pyrex frit (R) sealed to its bottom. Stirring is accomplished by immersing the end of the tube in the melt. To maintain a constant pressure under flow the input and evacuation rates must be equalized. To obtain this condition stopcock (C) is closed and flowmeters (B) are matched. The reaction is "stopped" by closing the inlet flowmeter and quickly evacuating the system. To obtain a liquid metal sample, the system is evacuated with the sampler positioned near the surface of the liquid. The sampler, Fig. 2, is then immersed until its thermocouple reading matches thermocouple (G) Fig. 1. The system is then pressurized with helium which forces the solution through the frit. Twenty to 30 min elapsed between the time the system was evacuated and sampled. During evacuation and pressurization of the system, the temperature variations did not exceed ± 5°C. While bubbling occurred the temperature remained constant to within 1/2 oC. The volume of the equipment was determined by adding a known volume of air at atmospheric pressure to the evacuated system and noting the resulting pressure.

Jan 1, 1962

By George K. Burke

Single axis soil mixing is a mechanical method of blending soil in situ with a reagent binder to achieve improved character. Generally this achieves a design compressive strength or shear strength and/or permeability, but can be used to immobilize and/or fixate contaminants as well as a treatment system for chemical reduction to a more "friendly" substrate. The reagent binder can be a variety of materials, including cement, fly ash, slag, lime, and an assortment of additives. Usually this reagent is delivered in a slurry form (combined with water), but the Dry Mix Method uses pneumatic conveyance.

Jan 1, 2001

By Richard Brummer, John Hadjigeorgiou, Patrick Andrieux

Some of the major challenges associated with underground mining at depth and/or under high extraction ratios are related to the resulting high stress levels, which can lead to the failure and collapse of the rock mass surrounding the excavations. Sudden and violent failures, which can have severe consequences, represent the highest degree of difficulty in terms of underground personnel exposure, erratic and unpredictable production, ground rehabilitation requirements, possible reserve losses and overall mining costs.

Jan 1, 2012

In most surface mining operations, overall project economics is driven to a significant extent by the slope angles that can be achieved within the rock masses present. This is particularly true of precious metal pits in which stripping ratios can be relatively high, although examples from other non-precious metal operations (such as large iron ore and copper pits) can also be found. As a consequence, it is not surprising that in most large (deep) open pit mines, relatively small changes in slope angle (of the order of two or three degrees can translate into changes in project value measured in the tens or even hundreds of millions of dollars. Understanding how this potential value can be realised, and quantifying the associated risks, should be one of the key outcomes of modern slope engineering practice. In striving for optimised slope angles, therefore, mining companies and their specialist consultants are increasingly turning to more sophisticated methods of engineering open pit walls. This paper summarises contemporary slope engineering practice as well as some of the key technical issues that are either currently being researched or will need to be researched in the future. It is written in the context of the slope engineering process; namely, data collection and analysis, failure mode identification and stability assessment, slope design, slope formation, and slope (face) mapping and monitoring (the latter two activities being the primary on-going data collection functions that close the design loop).

Jan 1, 2002

By Ray H. S, Galgali R. K

It is well accepted that the blast furnace will continue as the main route for iron-making for many more years. However, its limitations have motivated researchers to look for alternative routes, particularly, those where non-coking coal is used for liquid iron production. Smelting reduction (SR), where coal and oxygen are used to produce liquid iron and power is now receiving serious attention. Several smelting reduction reactors are now operating in pilot plant scale and one or two have reached commercial status.

Jan 1, 1999

By P. Somasundaran, D. W. Fuerstenau

This article presents the results of an analysis of grinding of 1:1 mixtures of 4x8-mesh quartz and limestone in laboratory ball and rod mills. From these experiments, the amount of energy consumed in grinding each component of the mixture was determined. In the ball mill, limestone consumes slightly more energy than quartz in the initial stages of grinding, but as the grinding proceeds quartz begins to consume a slightly greater fraction of the energy. The postulate that the fraction of energy consumed by any material in a ball mill is proportional to its volume fraction has been shown to be approximately true. On the other hand, in a rod mill energy consumption by quartz and limestone is equal only in the initial stages of grinding, with energy consumption by quartz greatly exceeding that by limestone as rod milling proceeds. Preferential energy consumption in tumbling mills has been discussed in a number of papers. For example, from an analysis of the kinetics of comminu- tion in ball mills, Arbiter and Bhrany' obtained rate equations based on the postulate that only coarse particles are ground, indicating that fine particles are protected from the action of the balls by coarser particles. In rod mills this protective action is much more pronounced than in ball mills. Recent studies on the rod milling of limestone-quartz mixtures have shown that this protective action causes the mineral with the lowest grindability (quartz in this instance) to consume the greater fraction of energy. 2 However, in the ball milling of mixtures it has been postulated that the consumption of energy by one of the constituents is determined by its volume fraction within the mill. 3 Possibly this hypothesis may be a simplification of the ball milling of mixtures since the protective action suggested by Arbiter and Bhrany may result in some preferential energy consumption by one of the constituents of a mixture. The objective of this study has been to determine preferential energy consumption in laboratory ball and rod mills by analyzing size distributions of each material in the products obtained from grinding 1:l mixtures of quartz and limestone. MATERIALS AND METHOD The procedure for carrying out the tests on wet grinding quartz-limestone mixtures has already been described.'r3 In the present investigation, 1:l mix-

Jan 1, 1963