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By Wallace N. Thomas

The present interest in large diameter grinding mills and the rapidly rising cost of the drive in proportion to the cost of the mill establishes a need to explore the drive system in detail. The drive for a seven to eight thousand horsepower mill will cost from 1350 thousand to $600 thousand which represents nearly half the total cost of the mill package, and requires as much building space as the mill proper. Some of the drive components require a longer time to manufacture and the installation time and costs are equal to those for the mill. It is apparent that the drive has become a dominant factor in the overall mill design consideration. A new plant requires a tremendous investment in both time and money. With the high degree of competition in the mining industry, low capital cost is an exceedingly important factor in the development of a project. It is just as important, however, to insure that the equipment is going to perform satisfactorily for the anticipated life of the project without adding unplanned operating, maintenance and replacement costs. This is the position in which the engineer or designer usually finds himself and is really one of his primary functions -- establishing a realistic compromise between cost and performance.

Jan 1, 1969

By T. Petrov

This study was performed to analyze the behavior of the blowing curtain face ventilation system modified with a specially designed regulator, called Wing Regulator. The purpose of the regulator is to extend the airflow penetration depth toward the face and allow effective face ventilation at curtain setback up to 13.7m (45ft) from the face. A small scale model of the Wing Regulator was successfully tested in the Mine Ventilation Laboratory at the University of Kentucky. Then a prototype of the proposed regulator was tested in full-scale in mine environment. Series of three dimensional simulations were performed to support this research using validated CFD code. This paper presents results of the performed studies and discuses the advantages of using the Wing Regulator. Presented information relates to the US Provisional Patent Application No 61/818,112.

Feb 23, 2014

By Doug Harding

The 3.5 km long Ashbridges Bay Outfall in Toronto, Ontario, Canada was a challenging drive set below Lake Ontario. After a remote machine acceptance due to the global pandemic, an 8 m diameter Single Shield machine launched in March 2021 from an 85 m deep shaft and began its bore in shale with limestone, siltstone and sandstone. During excavation, the TBM and its crew bored a city-wide record of 30 rings in one day, or 47 m of advance. This paper will cover the unique project, from TBM acceptance through to launch, tunneling in difficult conditions, and completion in 2022.

Jun 13, 2023

By Jan Van Greunen, Mark Vanderzee, David Haug, Geoffrey Hughes, Yiming Sun, Roozbeh Geraili Mikola

"The Sanitation Districts of Los Angeles County is planning to construct a new, approximately 7-mile-long (11.3 km) effluent outfall tunnel to be excavated by TBM and lined with precast concrete segments. Anticipated net operating pressures within the tunnel are expected to approach 3 bar. To counteract joint opening and exfiltration leakage, an approach has been investigated to post-tension the precast concrete tunnel lining with steel tendons. The tunnel alignment crosses several seismically active fault zones, at which installation of steel pipe is proposed as a secondary inner lining to mitigate potential offset displacements. This paper discusses the tunnel lining design approach and methods used to address these unique challenges.INTRODUCTION The Sanitation Districts of Los Angeles County (Sanitation Districts) is planning to construct a new 18-foot-diameter (5.5 m), approximately 7-mile-long (11.3 km) tunnel from the Joint Water Pollution Control Plant (JWPCP), located in the City of Carson, to the existing White Point outfall manifold structure at Royal Palms State Beach (RPSB) (Figure 1). The project is part of the Sanitation Districts’ Clearwater Program, which will improve overall system reliability through the repair and replacement of the existing aging infrastructure. The existing system includes two 6-mile-long (9.7 km) tunnels, 8- and 12-feet in diameter (2.4 and 3.7 m), and four ocean outfall pipes with a maximum capacity of approximately 675 million gal/day (MGD) (2.6 billion L/day). Addition of the new parallel tunnel will increase overall capacity to accommodate peak wet weather wastewater flows of up to 1,262 MGD (4.8 billion L/day) and will allow inspection and repair of the existing tunnels. The new outfall tunnel will be excavated using a pressurized face tunnel boring machine (TBM). Shafts at the north (JWPCP Shaft) and south (RPSB Shaft) ends of the proposed alignment will connect the tunnel to the existing wastewater treatment plant and existing ocean outfalls manifold and during construction act as launch and receiving shafts for the TBM. The tunnel will be excavated from north to south as a single heading."

Jan 1, 2016

By Andres Salazar, Carlos Delgado Robles, Richard K. Preece

The paper discusses the evolution of the geometallurgical model at the Escondida mine in northern Chile, the world’s largest copper producer. It highlights the transition from early metallurgical forecasting methods to a comprehensive model that includes spatial distribution of metallurgical forecasts. This model helps in planning and adapting to the variability in ore characteristics, such as hardness and mineralogy, which affect processing performance.

Dec 1, 2023

By Vas P. Srinivasa

A comminution circuit is an important integral part of a mineral processing plant. Nearly 40% of the capital cost and 30 to 40% of the operating cost are taken up by the crushing and grinding circuit. Therefore, considerable care should be given to the design and selection of this circuit. Several factors affect the selection of a comminution circuit--energy consumption, avail- ability of the equipment, the remoteness of location, and availability of trained personnel, to mention a few. Each of these factors influences the choice. Selecting a comminution circuit strictly on lower capital cost may not be optimal. Crushing and grinding circuits are usually site specific and should be selected - only after careful analysis of all the factors involved. The objective of this paper is to show the effect of these factors, which are extranious to capital cost. Usually, the capital cost of a semiautogenous grinding circuit is lower than that of an equivalent conventional circuit. However, due to lower equipment availability and high power costs in certain locations, the economic advantage of lower capital cost is lost. Analysis indicates that a break even point occurs at a power cost of 5¢/kWh. This paper describes the procedure utilized for selecting the most economic comminution circuit and states the conditions applicable for each cominution circuit.

Jan 1, 1983

By K J. Kidd

There is general understanding of hoisting systems throughout the mining industry. However, the relative importance of the various parameters influencing hoist selection and utilization is not completely understood. In order to determine if these parameters can be identified and prioritized, data covering 613 hoist installations were compiled. The 613 installations represent hoisting systems in 30 countries on six continents. This study compares data to historical trends, correlates design parameters in an attempt to identify underlying selection criteria, and comments on the apparent trends in recent developments in hoist design.

Jan 1, 1998

By R. J. Sawyer

The Greens Creek underground polymetallic mine restarted production in 1996 from a newly discovered high grade ore zone. Efforts to improve payable gold recovery have focused on gravity separation. The paper compares testwork and plant results for a newly installed gravity circuit.

Jan 1, 1997

By Alessandro Boscaro, Daniele Martinelli, Cristina Gabriela Oñate Salazar, Daniele Peila, Carmine Todaro

"The wear phenomenon in TBM machines is not a simple issue, as many parameters are playing an important role on the consumption of the tools during a tunnel excavation. The action of the wear due to the contact with the soil can occur on the excavation tools, on the cutterhead, in the excavating chamber and on the screw conveyor. The test developed at the Laboratory “Tunnelling and Underground Space Center” of the Politecnico di Torino is aimed to reproduce the interaction between the tools and the excavated material, simulating the excavation conditions in terms of soil conditioning and water content. The test, which is not standardized, is studied to evaluate the wear in a global point of view. The present work starts by testing the influence of the use of chemical agents during the conditioning phase on the drop of the wear phenomenon, in terms of metal weight loss. 1. INTRODUCTION The wear phenomenon of the excavation tools used in TBM tunneling machines is playing recently a major role in the design stage of full-section tunnels in both soils and rocks. Concerning EPB excavation technology, the main issue regarding productivity is the wear of its components (Peila et al.,2012), which increases the so called dead time in the whole process. The friction generated between the components and the ground causes the consumption of the tools and other metallic components in contact with the ground, described as primary and secondary wear by Nielsen et al. (2006a). This leads to downtime, with risky and complicated maintenance operations (Peila and Schulkins, 2009), as occurred in recent tunneling works in Liguria (Italy). The study of this argument in granular soils is very delicate, as it encloses not only the properties of the excavation tools, but also the characteristics of the ground to be excavated (hardness, mineralogy composition, shape and dimension of the particles and moisture content). The penetration mechanism is based on the scraping of the ground, by mean of cutting tools, with its consequent consumption. Concerning the rocks, the excavation principle is characterized by crushing the rock mass matrix by applying high compressive stresses through cutting disks. In such a way, to avoid frequent downtimes and their consequent economic losses, it is very important to study this subject and therefore understand and forecast the durability of the components, in order to try to maximize the productivity of the machine by reducing the frequency of maintenance stoppings."

Jan 1, 2016

By John J. Dran

Large mining companies use a number of sophisticated analytical techniques for introducing the element of risk into the valuation of mineral reserve sites. These methods include, but are not limited to, sensitivity analysis, computer simulation, conditional probability, and portfolio theory. Such techniques enable the managements of these large companies to have an awareness of the many possible ramifications of their investment decisions. On the other hand, small mining companies and individuals in the mining business, due to limited financial and technical resources, still rely primarily on single point estimates of the value of mineral reserves. These small miners introduce the aspect of risk only through the application of an appropriate "speculative" discount rate. This paper illustrates the application of a technique for the introduction of risk into investment analysis. The method appears to be most suitable for use by small mining companies and individuals contemplating a large investment in a new mineral reserve site.

Jan 1, 1974

By F. Stanley, A. Noble, G. Luttrell

"The supply chain for coal-fired power stations involve a multi-stage network of mining units, preparation plants, blending facilities and transportation systems. Attempts to optimize the supply chain have historically been difficult due to the lack of communication/coordination between the various groups involved in coal production, transportation and utilization. In light of this issue, a new approach for globally optimizing the coal supply chain called Micro-Price Optimization has been developed to simplify the optimization problem and to provide coal producers with fundamental insight needed to improve the profitability of their operations. This fundamental concept assigns unit values to each particles (or groups of particles) passing through production units. Optimization occurs whenever the maximum tonnage of positive value particles are recovered and passed to market, while the minimum tonnage of negative value particles are recovered. Elements considered by the concept include factors such as target boiler cost/efficiency, waste disposal costs, coal quality parameters, emission penalties, transportation charges, or other quantifiable impact factors. This article reviews the basis for the micro-price concept and provides a detailed industrial case study involving the determination of optimum coal qualities that need to be supplied by mine operators to feed a coal-fired power station. INTRODUCTION The performance of coal preparation plants have long been known to play an important role in determining the overall profitability of mining operations. This significance is due to the fact that the misplacement of just 1 ton/hr of clean coal to reject resulting from poor plant performance can represent nearly $300,000 per year in lost revenues for a mine producing a steam market coal (i.e., 1 ton/hr x $50/ton x 6,000 hr/yr = $300,000/yr). This loss can be even more pronounced, occasionally approaching $1 million per year for each 1 ton/hr of misplaced clean coal, for operations producing high-quality coking coal products sold into the premium metallurgical market (i.e., 1 ton/hr x $150/ton x 6,000 hr/yr = $900,000/yr). Detailed economic studies indicate that improvements in plant yield resulting from increases in organic efficiency have one of the largest overall impacts on mine revenue when compared to other cost centers within the mining operation. One of the first studies to show this important financial impact was reported by Norton-Hambleton Consultants (Norton, 1979). Their study indicated that an unfavorable change in plant efficiency was second only to coal transportation in terms of its impact on profitability, while an unfavorable change in the cost of underground mining was a distance third. This longstanding realization has continued to provide a strong financial incentive for coal producers to optimize the separation efficiency of their coal processing and blending operations."

Jan 1, 2015

By W. W. McClintock

A column flotation system was designed to conduct release analyses of a very finely disseminated, stratiform, copper ore from the Denali Copper Prospect, Alaska. Chalcopyrite is finely disseminated in a carbonate matrix and pyrite is disseminated in both the chalcopyrite and gangue, requiring grinding to 20 microns for adequate liberation. A 25 mm x 2.4 m glass flotation column was designed. The ore and its flotation tailings were repeatedly floated to obtain a rougher concentrate, which was repeatedly floated to produce release fractions. These were analyzed for copper, iron calcium and magnesium. Release curves were plotted. The procedure was replicated at each of two grinds; 99% and 88% passing 37 microns (400 mesh), and should find application for flotation characterization and process selection. Bond grindability work was also conducted.

Jan 1, 1994

Contents

Jan 1, 1985

By J. Alderman

In modern industry, cost and productivity concerns and environmental concerns are sometimes incompatible, but at other times only appear to be incompatible. Surface disposal of wastes from coal preparation plants appears to be an area of the production system with little opportunity for initiating measures which cut costs and concurrently benefit the environment. However, improving the recovery efficiency of fine coal circuitry can cost-effectively reduce the plant waste environmental burden by a substantial amount. Studies at Jim Walter Resources indicate that fine-coal waste measuring 50% ash contains over 50% recoverable, marketable coal. Modifications to the plant circuitry which reduce the amount of misplaced coal in the fine-coal waste stream could substantially reduce the volume of fine waste disposed aboveground. Studies at Jim Walter Resources indicated that episodic overloads of the vacuum filters in the fine-coal circuits at the preparation plants were a source of misplaced coal in the fine-waste stream of the cleaning plants. To improve the filtration rate and reduce overload episodes, conventional rotary blowers were replaced with high pressure air blasters on most of the filters. Subsequent testing indicated that measurable benefits can be derived from this modification.

Jan 1, 1990

By R. Schipper

Two bench scale biochemical reactors (BCRs) were operated over a period of 3 months for the removal of uranium, radium, and selenium from mining influenced water (MIW). A BCR is an anaerobic reactor comprised of organic and inorganic substrate including woodchips, straw, sawdust, and limestone. Uranium and selenium are removed in a BCR by biological reduction of oxidized species (i.e., selenite, selenate) to less soluble forms (Pahler et al. 2007, Blumenstein et al. 2008, Rutkowski et al. 2013). It is believed that radium may be removed through absorption to the biomass in the BCR. The bench scale BCRs were constructed of 6 inch PVC columns with a top to bottom flow configuration. The bench influent water consisted of three different water ratios from two adit seeps. BCR hydraulic retention times of 24 to 36 hours were targeted during the study. Influent concentrations ranged from 19 ug/L to 1.86 mg/L for dissolved uranium, 12 ug/L to 270 ug/L for dissolved selenium, and 4.9 pCi/L to 24 pCi/L for radium 226. The average and maximum dissolved uranium removal was 93.6% and 98.5% respectively, the average and maximum dissolved selenium removal was 93.3% and 98.9% respectively, and the average and maximum radium 226 removal was 89.5% and 100% respectively.

Feb 23, 2014

By J. Marbeau

In this paper, a geostatistical study of the North Butte roll type uranium deposit is presented. The geostatistical procedure applied takes into account the varying direction of the roll front for reserves calculations. For variogram calculation, the roll front is unfolded thereby allowing estimation of the variogram in directions parallel to and perpendicular to the roll front. Variograms calculated in unfolded space show strong directional anisotropies consistent with the anticipated configuration of the deposit.

Jan 1, 1991

By N. D. Cristescu

The aim of the present paper is to show how a mathematical modeling of a creeping rock and that of an yieldable lining can allow us to obtain an optimization of the lining design and generally of the whole procedure and sequence of excavation-lining application. We follow the procedure presented by Cristescu (1989) Chap. 13, and Cristescu et al.(1987); for some other approaches see the literature mentioned, for instance, by Ladanyi and Gill (1988). We have in mind to describe the creep of the rock, i.e. the tendency of the tunnel to close in time, and to take into account that the lining itself may be deforming, i.e. possesses some rheological behavior. We start from the well known fact that "intelligent" lining, which are yielding to the pressure of the rock, may be much more successful in supporting creeping rocks than excessively strong, badly designed supports which are unable to resist to pressure of the rock, though these are oversized supports (too strong, too stiff, installed too early, etc.). In order to optimize the lining to be used for a certain rock, it is absolutely necessary to study in the laboratory the main mechanical properties of that specific rock and to formulate for it a reasonably simple mathematical model. Standard triaxial test results are quite enough to determine the model for the rock, combined with some creep tests which have to reveal the long-term behavior of rock (the so-called rheological model). What concerns the lining, we consider several kinds of linings existing today, for which a mathematical model was also formulated. One shows which is the interaction between the creeping rock? and the yielding lining and how can it be formulated in mathematical terms. Once written in mathematical terms, an optimization of the lining design and that of the excavation layout and support application sequence is possible. We will try to make the exposure as simple as possible, avoiding details and having in mind to give only a sketch of the problem, while the mathematical details could be worked out quite easy.

Jan 1, 1994

By J. D. McKay

Continuous column flotation parameter testing, to improve operating methodologies, using coarse bubbles on fluorite ore from Nevada, revealed that retention time is the predominant factor affecting column flotation fluorite recoveries. Column height and feed injection location affected fluorite grades and recoveries because they affected particle retention times in the collection and cleaning zones. Other parameters tested, including froth depth, wash water addition, and axial mixing, showed varying results. Four defined zones within the column were found to have characteristic rates of upgrading per vertical foot. Fluorite-gangue separation was more efficient in a column than in a conventional cell at all particle sizes.

Jan 1, 1987

By Z. Yanni, W. Xuan, Z. Jingyu, D. Jun

"In order to conserve the coal resources, many coal mines which have been mined begin to re-mine. Re-mining face is of complicated circumstances and coal spontaneous combustion is unknown, so experiments are designed to study the characteristics of multiple oxidized coal samples. The investigation into the oxidation of coal is conducive to the safe mining of the coal seam groups and thick coal seams. The non-caking coal is collected as coal samples. The raw coal sample with initial oxidation and the sample after second oxidation and third oxidation are obtained after a certain treatment of the coal. The temperature-programmed system, and proximate and ultimate analysis are adopted to compare the releases of gaseous products of spontaneous combustion characteristic parameters of the multiple oxidized coal sample. With the increase of oxidation times, experimental results indicate that the oxygen consumption rate, the release rate of carbon monoxide and the intensity of heat release are significantly decrease. The characteristic temperature dropped and the size of the pores in coal is smaller as well. INTRODUCTION China accounts for 13% of global coal resources [1-2]. In recent years, more and more coal mines begin to re-mine in order to conserve resources. This production mode is to mine coal pillar left by last exploitation or unavailable for mining. In the process of slicing mining for coal seam groups or thick coal seams, the second or multiple oxidation due to the mining-induced effect, may take place in the goaf of the lower seam. So the spontaneous combustion is prone to occur in this area. The study on oxidation of coal is conducive to the safe mining of the coal seam groups and thick coal seams. Scholars [3-7] have discussed the workability of re-mining for coal mines, verify the feasibility of the existing mining technology for re-mining area, and set up the prevention measures of coal spontaneous combustion. But all these measures are only based on on-site observation data, instead of experiment on spontaneous combustion characteristic parameters of re-mining coal samples which have a great importance for early prediction and forecast. The authors aim at investigating the multiple-oxidated coal spontaneous combustion law, and perfecting the coal spontaneous combustion mechanism to provide a basis for the prediction and forecast."

Jan 1, 2015

By E. J. Duggan

In the course of the grinding experiments conducted at Climax over the past twenty years one interesting phenomenon was discovered. This was that balls made of certain types of steel grind more ore than some others. The most significant variation was found in the results secured with a hard alloyed forged bail compared with the forged carbon ball supplied by most ball manufacturers These findings were confirmed in many repeated controlled tests. The value of these tests depends on the methods used, the facilities available and the care with which the investigations were conducted. These are described in the following. This testing has largely been done in eight 91 x 81 low discharge ball mills in closed circuit with 7811 duplex screw classifiers. The mills run at 20 RPM on a feed approximately minus 3/8 inch. The ore is ground through 28 - 35 mesh in one stage. Each mill and auxiliary classifier is in a separate, parallel grinding and flotation section The, ore to each: section is the same in every particular and is fed to each mill bin from a common conveyor tripper. The ore from the bin to the ball mill is weighed on a conveyor scale. The power input to each mill motor is determined with a watt hour meter which measures the power to the mill motor only. Both the scales and meters are checked periodically by standard methods and have been found to give consistent readings. Automatic samplers are employed on each section to obtain accurate measurement of the size reduction. A size analysis of the classifier overflow is made each shift and a composite size analysis is obtained from a weighted average of the various shift analyses. The mills are driven by 450 horsepower synchronous motors so that there is no speed variation and shell liner age is kept as nearly the same as possible in mills under test. The same set of operators is assigned to a mill under test as to the control mill. In all other particulars such as ball load, mill density, liner design, etc., the mills are operated in precisely the same manner. This is an ideal arrangement for plant testing. From observations of ball mill performance over the past twenty years, it is the opinion of the operating staff that variations of as little as one to two percent in mill capacities may be determined if sufficient operating hours are devoted, to a test. Relations between mill capacities, fineness of grinds and power consumptions have been established with a high degree of accuracy. In Table I are given the figures for one test. These show that while the grind with the hard ball was slightly coarser, the tons ground per hour was 5.4 percent greater with the soft ball, and 6.1 percent greater on the basis of minus 100 mesh produced. The power consumed per ton with the hard ball was 4.4 percent greater on a tons per hour basis. The soft ball was inferior in wearing qualities and steel consumption per ton of ore was 24 percent greater with it. In this test the mills were charged with the two different balls for a period of three months to allow the ball loads to come to equilibrium. In Table II are given the specifications of the two balls. The difference in grinding rates may be accounted for by the nature of the ball surfaces. The surfaces of the soft balls after being worn a while are rougher to the touch. Enlarged photographs of the worn-in surfaces of hard and soft balls are shown in Figure I. These show the presence of scratches or skid marks on the hard ball surface which indicates less nipping effect. The Climax ore pulp has a pH of 6.0 and is in a dispersed state. The viscosity is low compared with most ores especially those that are conditioned with lime. This would lower the tendency of the pulp to adhere to the balls with a lower grinding efficiency as a result. It was also thought that the relative rates of wear from surface to center might affect the make-up of an equilibrium ball charge. This probably has little effect on capacity as shown by the size analyses of the two ball charges in Table III. This hypothesis was subsequently investigated by the use of rationed charges. The rationed charge did not overcome the advantage in

Jan 1, 1958