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By P M. Outhwaite

Everyone present today is especially aware of the need for minerals in this society of ours. The words 'especially aware' were deliberately chosen because this is a meeting of the NZ Branch of The Australasian Institute of Mining and Metallurgy not a dilettante group such as some of our critics. Members of the Australasian Institute of Mining and Metallurgy knew that our great farming industry, our shelter, our transport and our energy supply, are largely dependent on a flow of minerals into this country despite the presence of large coal and natural gas reserves. Flow in was said because New Zealand imports most of its mineral requirements with all the consequential strategic and economic disadvantages. If the above statements are accepted by this gathering it would be a premise fairly stated that New Zealand has a great need to discover and develop minerals.

Jan 1, 1976

By Antonio Nieto

According to the analysis of surface powered haulage accidents involving off-highway haul trucks in openpit mining, the U.S. Mine Safety and Health Administration (MSHA), reported 1,300 fatal accidents from 1990 to 2005. Of these, 163 were considered off-highway truck haulage accidents, and at least 23 involved collisions of an off-highway truck with a neighboring vehicle or a worker on foot (Nieto and Dagdelen, 2006). In general, an average of 18 nonfatal accidents and three fatal accidents occurred in openpit mining per year when equipment backed over the edge of an embankment, stockpile or dump point (Fessak, et al., 2003). The specific problems that need to be solved can be determined by carefully studying the accidents involving haulage trucks (Marrara, et al., 2004). Solutions to haulage truck safety must consider the human factors affecting each task even when engineering solutions seem most appropriate. In surface haulage, human performance becomes a critical issue because of the unusual demand the vehicle?s places on their human operators: ? Roadways and working areas change frequently in mining operations. ? The sheer mass of mining equipment such as trucks requires control inputs (e.g., braking) far in advance of the desired action. ? Rough roads, loading and unloading impacts can subject the driver to dangerous shocks and vibration. ? Visibility is sharply curtailed by the bulk of the vehicle generating large blind spots. All of the above issues can be attributed partially to the lack of visibility (Pyles et al., 2003). Driver visibility and communications are key issues to reduce mining truck accidents. Analysis of fatalities and injuries involving mining equipment demonstrates the need to develop new intervention and control strategies to mitigate the risks associated with mining equipment operation. Ongoing development of these and other new strategies is required to achieve a reduction of injury in the mining industry

Jan 2, 2010

By Walter E. Vanderpool

Three drilled foundations for an elevated water-tank were instrumented, tested and monitored during construction and the first two years in service. One shaft was load tested by the Osterberg method. A void was detected in one shaft by cross-hole sonic logging, pulse-echo logging and concrete coring. The void was instrumented, hydrostatically tested and pressure grouted. The shaft was ?abandoned? and replaced by two smaller shafts with a straddle beam cap. One of the replacement shafts and the cap were instrumented. The strain that appeared in the abandoned shaft, the replacement shaft and the cap were monitored during the remainder of the construction, through the commissioning test and the first two years of service. The results demonstrate the precision and sensitivity of the vibrating wire strain gage.

Jan 1, 2004

By L. B. Hales

Expert control of grinding and flotation plants has been successfully used in the minerals industry since the 1970?s. The earliest of these systems were written in a hard-coded fashion in FORTRAN, BASIC or Pascal. Second generation systems were built using the first experimental expert system shells that were being developed in the artificial intelligence community. Later systems were deployed in expert system shells designed for real-time processing plants that also included the ability to model the process with neural network models and optimize setpoint selection through the use of genetic algorithms. Significant performance increases have been achieved using these systems, but in general, they suffer from the static nature of their rules and to a degree the process models also. Typical improvements quoted in the literature suggest throughput rate increases of 4 to 8 percent and recovery increases of 1 to 4 percentage points. In over 30 years of designing and installing expert systems we have never seen performance numbers less that these but have often seen numbers in excess of the high values. However, once the system has been installed and tuned up the performance usually does not improve much thereafter. Granted, the initial increases in throughput rate and recovery are significant but a question always lingers, that is, can the expert strategies be modified further to squeeze a little more improvement out of the system. The answer to this question is a resounding yes, however a complete understanding of what we do well as designers and operators of expert control systems is required as well as how we can improve upon our best practices is required to move us from the performance plateau we have been on for a number of years to a new higher level of performance.

Jan 1, 2009

Lonmin, earlier than any other primary platinum producer, had to deal with the concentrating and smelting of UG2-rich ores and concentrates respectively. Smelting was performed at a fairly modest scale compared to the company's industry peers in the platinum group metals (PGMs) industry, and was focused on smelting concentrates obtained through low mass pulls while still maintaining high recoveries at its concentrators. Lonmin gradually smelted larger quantities of UG2 concentrates. Initially two 2.3 MW Infurnco furnaces were commissioned in 1982, followed by three 5 MW circular 3-electrode Pyromet furnaces in 1991. Deep electrode immersions and moderately high hearth power densities were used in all the designs. Lonmin decided to continue with circular furnace technology when it planned its new high-intensity No. 1 Furnace. Neither Lonmin nor the technology supplier and EPCM company foresaw the challenges that scale-up would bring when applied to the smelting of UG2-rich concentrate blends. Superimposed on the high chromite content, was the low base-metal loading, mineralogical difficulties, and fine particle size that resulted from milling of all concentrates (especially UG2 and recycle materials within the smelter). Through a process of fundamental diagnoses of furnace run-outs and wear patterns observed during repairs, the main challenges operating this high-intensity furnace were identified as sulphur vapour corrosion of the copper coolers, uneven and unpredicted movements of refractory bricks with associated copper cooler lift and matte tapblock movement, the formation of three-phase 'mushy' zones, high refractory and taphole wear rates, uncertainty in matte level and associated insufficient matte buffer height, and high furnace operating temperatures. These factors seldom worked in isolation and will be explored in more detail in the paper. The variability in furnace feed characteristics led Lonmin to redesign Furnace No. 1, to invest in backup furnace capacity, and to invest in improved monitoring and control. These improvements consisted of online and high-frequency off-line monitoring of feed chemistry and mineralogy, online pressure monitoring of water-cooled circuits, an alternative matte liquid level measurement in the furnace, and electrode immersion estimation. Improvements were also made at the converters by installing and utilizing optical spectral analysis of the converter flame to characterize converter behaviour and achieve the desired iron end-point for white matte. This paper also briefly discusses the current and future expansion plans, as well as ancillary operations at the smelter such as flue-gas handling and materials handling.

Jan 1, 2011

By L. "Andy" Anderson, Stephen N. Short, James E. Friant

Since the introduction of "boxhole" drilling in South Africa in 1973, the concept of a blind raise, drilled upward from below, has become an accepted mining practice. A combination machine, capable of constructing raises to a mining level below, or blind raising upward from a single setup position adds a further dimension in versatility. A machine with these dual operating modes was designated by The Robbins Company as the 53R and was introduced into the Western Deep Level Mines, South Africa in 1984. The unit is powerful (200 hp) and compact (18,000 lbs., in a 10 ft x 4 ft 5 inch x 4 ft 6 inch package). This paper describes the design considerations for the new machine, provides field data from early trials, and looks briefly at its overall market and potential.

Jan 1, 1985

The ore, as sent to the sulphide mill, is a mixture, in varying proportions, of oxidized ore from above the 450-ft. level and sulphide ore from above and below (to date mainly from above) that level. The oxidized ore consists essentially of quartz, with varying amounts of kaolin and clayey material which passes into the slime and, in excess, creates a serious problem in settling. Pyrite may be present in small amounts. The gold is free and extremely fine.The ore, classed as sulphide, consists of pyritic material in a quartz gangue. The sulphides have been exposed to weathering, leaching by descending solutions and attack by ascending gases and actual burning. The alteration products occur in widely varying proportions and this variation adds to the difficulties in settling and control of grinding and flotation.Copper Minerals.-Chalcopyrite is the principal copperbearing mineral with smaller and varying amounts of covellite, which is sometimes associated with chalcocite, both these minerals being apparently of secondary origin.Examination of tailing by Dr. Stillwell has revealed all stages between clean chalcopyrite, chalcopyrite tinged with covellite, covellite with residual cores of chalcopyrite and clean covellite.The chalcopyrite is free floating, very few particles of clean chalcopyrite being observed in the tailing from granular flotation. The chief tailing loss is probably as chalcopyrite occurring as small inclusions in pyrite...

Jan 1, 1939

Similarities and Dissimilarities in Florida Phosphate Ore Types - Surface and Bulk Property Analysis of the Flotation Feed, Concentrate and Tails

Sep 13, 2010

By M. J. Sapko

A moderately scaled apparatus was developed by the Bureau of Mines to determine the fire resistance characteristics of mine conveyor belts and similar type materials. The design of the apparatus was based upon data obtained in full-scale fire tests and features a radiant panel to preheat the belt and a methane-oxygen ribbon burner to ignite the sample under ventilating condi¬tions in a rectangular duct. The test method overcomes the limitations of existing laboratory-scale methods and provides a measure of both ignitability and flammability in quantitative terms. Flame propagation depended upon such variables as air velocity, ignitor heat flux, and belt width and height in the test chamber. Data are presented for nine belt materials and fire resistance ratings are proposed in terms of the flame spread rate, heat release rate, and the critical ignitor heat flux. The ratings discriminate between the fire hazard posed by different fire-resistant belts and tend to be in reasonable agreement with those obtained in full-scale fire tests.

Jan 1, 1981

By Antoni Kidybinski

Mechanical properties of the nether roof rocks play a substantial role in maintenance of mine openings. On longwall faces too low strength is a cause of rock- falls which bring about delays in face advance, and hazard to the miners, while too high strength may result in over hanging of roof rocks in gobs, sometimes over a large area, initiating dynamic periodical pressures. Hence, efforts have been undertaken to develop methods of caveability evaluation for proper selection of roof control system and of longwall-face supports. DESCRIPTIVE CLASSIFICATIONS One of the basic qualitative criteria for classification of roof types with different caveability is the criterion of geological structure. The 26-stages scale worked out for conditions of Donetz Coalfield. USSR, by E.T. Proyavkin et al.(l) may be an example of that type of classification. This scale is presented graphically in Fig. 1. The first types of roofs according to this scale are easily caving immediately after exposure of a small surface, while the last be- long to the strongest roofs which over-hang in waste and require artificial fracturing. Observations and measurements carried out also in Donetz mines lead W.T. Davidyanc (2) to following classification: Class I - weak and densely bedded rocks very easily caving after each advance of supports. The roof subsidence over a waste edge is: [ ] where: h is roof subsidence, m M is thickness of the seam being mined, m S is width of an opening, m Class II - medium caveability and small main roof movements [ ] Class III - hardly breakable rocks, thicker beds with higher strength: caving by large blocks. After exposure of larger roof surfaces the movements of the main roof strata may affect an opening. [ ] Class IV - very hardly breakable rocks - thick, monolithic sandstones: they cause considerable difficulties in working with caving. [ ] Class V - thin-bedded rocks featured by the capacity of a smooth subsidence without disintegration into loose blocks with keeping contact between the particular blocks in a bed. In this classification mining with full roof caving is recommended for class I-III and only exceptionally for class IV. In many coal fields, qualitative and quantitative observations of roof rocks behavior have been carried out: as a result relations and characteristics found in certain mines are given. As an example, Kizielov Coalfield data in the USSR given by B.S. Vinokur (3) may be mentioned. In working thin pitching seams man- less system of undercutting the seam by means of rope saw is used there. The author separated four classes of roofs and gave, for every class, the empirical relation between the extracted area and the time of roof stability. The size of the stable area of exposed roof for particular classes is as follows: 120 to 150 m2 (1291 to 1614 ft2 ), 170 to 200 m2 (1829 to 2152 ft2), 250 to 280 m2 (2690 to 3013 ft2 ) and 350 to 400 m2 (3766 to 4300 ft ). For Kuznetz coalfield in the USSR the following classification of roofs was developed at the WIMI Institute in Leningrad: Class I - completely unstable: impact strength in- dex of rock f = up to 1.5 - the strip of roof 1.4 to 2.0 m (4.59 to 6.56 it) wide, exposed after shearer's cut holds stable during the time not longer than 5 min. Class II - low stability: f = 1.5 to 2.5 - the strip of roof 1.4 to 2.0 m (4.59 to 6.56 ft) wide and up to 4.0 m (13 it) long, exposed after shearer's cut holds stable during 5 to 20 min. Class III - medium stability: f = 2.51 to 3.5 - the strip of roof 1.4 to 2.0 m (4.59 to 6.56 it) wide and 4.0 to 25 m (13 to 82 ft) long, exposed after shearer's cut holds stable from 20 min. to 2 hours. Class IV - stable: f = 3.51 to 4.5 - the strip of roof 1.4 to 2.0 m (4.59 to 6.56 ft) wide and 25 m (82 it) long exposed after shearer's cut is stable during the time longer than 2 hours. Class V - entirely stable: f = over 4.5

Jan 1, 1982

By Paul G. Swanson

This investigation was conducted to identify the principle causes of spoil instability problems in the Eastern Coal Province. The study was also conducted to develop a matrix to predict the strength and groundwater parameters necessary to conduct stability analyses on a proposed reclamation scheme. Suggestive methods to reduce spoil instability potential are also included. Numerous physical property, strength, compressibility, chemical and mineralogical tests were conducted on mine spoil samples obtained from 11 different surface mine sites. Site characteristics that influence the position of the post reclamation groundwater table were identified. Distinct regions within the Eastern Coal Province with high incidences of surface mine slides or high landslide susceptibility were identified. The application of the predictive matrices developed in this study is for surface mine sites mined and reclaimed in general accordance with the Surface Mine Reclamation Control Act of 1977. A review of current mining practices and the relative effect on spoil stability is included.

Jan 1, 1983

By Josh Johnson, Stephen Tadolini, Mark Ferster, Curt Jacobs

"Stillwater Mining Company is the only U.S. producer of platinum group metals (PGMs) and the largest producer outside of South Africa and Russia. The Company controls a considerable portion of the J-M Reef and is currently developing the ""Blitz Project,"" which will provide a main haulage level, ore pass systems, backfill plants, and flow-through ventilation to increase production and ensure sustainable extraction of this critical strategic mineral. While driving two sub-parallel footwall laterals to about 8,125 m (26,000 ft) long, one of the development drifts experienced a massive fall in a geologically disturbed area created by a dike intrusion. Combinations of a void fill material (Tekseal), polyurethane injection, fore poling coupled with steel arches, and supplemental bolting allow safe and efficient advance through this geologically disturbed zone. This paper presents the details of this project.INTRODUCTIONStillwater Mining Company is the only U.S. producer of platinum group metals (PGMs) and the largest producer outside of South Africa and Russia. The Company controls a considerable portion of the J-M Reef and is currently developing the ""Blitz Project,"" which will provide a main haulage level, ore pass systems, backfill plants, and flow-through ventilation to increase production and ensure sustainable extraction of this critical strategic mineral.Tunnel Boring Machines (TBMs), shown in Figure 1, have been used extensively at both the Stillwater Mine and the East Boulder Mine to develop access to the J-M Reef. The J-M Reef is a PGMbearing reef deposit located in the Stillwater Complex near NyeMontana. The mining property and location are shown in Figure 2.The Blitz Project is named after the Blitz mining claims and is developing the Eastern extent of the J-M Reef Project., This consists of a 5.5 m (18 ft) diameter TBM drive on the SOE (1,562 m (5,000 ft) above sea level and east of the shaft collar) and a 4.7 x 4.7 m (15 x 15 ft) drill and blast heading on the 5,600E (1,750 m (5,600 ft) above sea level and east of the shaft collar). The Benbow Decline will be completed running perpendicular to the TBM by drilling and blasting methods. The TBM drift will be used for exploration drilling, ventilation, and will serve as the main haulage level. The 56E will be used for ventilation and exploration drilling. The Benbow decline will serve as the exhaust drift."

Jan 1, 2016

By B. McCosker

The choice of the appropriate delivery process for a project is often not adequately thought about. The industry needs to be prepared to break its preconceived mind-sets on delivery procedures. Industry should also be prepared, on a project by project basis, to commit to a critical review as to how the assessment and then selection of the best procurement method can best optimize the project outcome for both client and contractor. Conventional standards and biased reliances and experienced based preferences should be challenged. To achieve this requires a paradigm shift in attitude to both the process of tendering and the evaluation of tenders. Tender and contracting practices are dynamic and have moved away from solely cost based assessments. Tendering contracts arise imposing changed duties upon owners. Traditional contract delivery models have shifted, reflected in part by the trend towards relationship contracting. The categories of contract options are not closed. Many are but hybrids of other time-tested models. This paper considers options available for any brownfield or greenfield project. The principles outlined are translatable to supply, upgrade, operation and maintenance arrangements.

Jan 1, 2002

By B. K. Gorain, A. W. Stradling

"The traditional metallurgical scale-up techniques for flotation plants are based on lumped flotation rate constants, dimensionless hydrodynamic numbers and the use of safety factors. Once the plant has been designed, the flotation principles that govern scale-up play almost no role in the optimization of the circuit to meet its metallurgical targets. This deficiency can potentially lead to a plant design incapable of meeting its targets, especially when circuits incorporate large flotation cells. This paper demonstrates the application of the k = PSbRf model in providing a suitable framework for flotation modelling, scale-up and circuit optimization, thus presenting an integrated approach to the problem. Studies were conducted at Teck Cominco’s Red Dog concentrator to characterize cell hydrodynamics, gas dispersion properties, froth flow behaviour, entrainment and ore floatability to obtain the key data necessary to develop a model for the flotation circuit. The findings of these studies are presented in this paper, which provide valuable insights into the modelling and optimization of the Red Dog flotation circuit. IntroductionThe traditional methodology for flotation scale-up and plant design is based on the ruleof- thumb with safety factors and dimensionless hydrodynamic group similitude considerations. With the recent trend toward larger flotation cells with volumes of 150 m3 and greater in concentrators, the problem of scale-up has become even more acute. The scale-up factors that may have been tolerated on smaller cells are not applicable to larger cells. The transportation of froth in these large cells adds further complexity to the problem. Arbiter et al. (1980) foresaw this problem and pointed out that if scale-up were attempted by rule-ofthumb and trial-and-error, the probability of failure would then increase with scale. This problem is confounded by the fact that the design and scale-up considerations used for a new flotation circuit play almost no role in optimization of the circuit to meet its metallurgical target. This results in a plant design, which may be incapable of meeting its metallurgical targets. There have been several instances where new concentrators have required significant circuit modifications and additional capacity to achieve design targets."

Jan 1, 2003

By D Craw, K A. Hunter

Deposits of the metalloids arsenic and antimony occur in quartz veins of mesothermal hydrothermalsystems in the South Island as arsenopyrite (AsFeS) and stibnite (Sb2S3). Arsenopyrite and stibnite occur in natural exposures in gold-mining areas, and additional arsenopyrite and stibnite have been exposed to the atmosphere by mining activity. At present the drinking water limits for dissolved As and Sb are 10 ppb and 3 ppb respectively. Sampling and analysis techniques are critical to determination of metalloid concentrations at these low levels. Endeavour Inlet in the Marlborough Sounds is an historic mining area (late 19th Century) with elevated As and Sb in waters emanating from mineralised rocks. Waters running out of abandoned adits have concentrations of As and/or Sb up to ppm levels. Natural waters have high Sb and low As, and the As levels increase markedly where anthropogenic activity has exposed rock. Changes in Sb/As ratio could potentially be used to quantify the anthropogenic component of metalloid mobility on mine sites on the 100-year time scale.

Jan 1, 2002

By R. R. Joseph, K. A. Gschneidner

The solid solubility of magnesium in the close-packed modifications of lanthanum, cerium, praseodymium, neodymium, gadolinium, dysprosium, and lutetium was determined from approximately 250°C to the eutectoid temperature by the X-ray parametric technique. The maximum solubility at the eutectoid temperature was found to be 9.4 at. pet in La, 8.2 in Ce, 9.2 in Pr, 8.2 in Nd, 14.1 in Gd, 13.5 in Dy, and 11.5 in Lu. The solubility when compared at a homologous temperature was found to increase in a smooth manner with decreasing size factor from lanthanum to lutetium. It was found that the apparent atomic radius of magnesium when dissolved in the lanthanide solvent was constant and could be accounted for by consideration of several models dealing with departures from Vegard's law. The eutectoid temperatures for the lanthanide-rich La-Mg, Ce-Mg, and Nd-Mg systems were found to be 544", 505°, and 551°c, respectively. A study of the solid solubility of magnesium in the lanthanide metals was initiated because the valency and electron egativity difference are essentially constant and favorable for solid-solution formation, and thus one could examine the effect of size on terminal solid solutions. The size factors based on metal radii range from 14.6 pet for lanthanum to 7.5 pet for lutetium. Since the electronegativity difference is small, 0.06 unit or less, application of the Hume-~othery' and Darken and Gurry2 criteria for solid-solution formation indicates that the solubility of magnesium would be expected to be small for lanthanum (between 5 and 10 at. pet) and to increase as we proceed along the lanthanide series. A search of the literature revealed no data on the solubility of magnesium in the lanthanide metals except for the solubility of magnesium in cerium at 450oCA typical phase diagram for the lanthanide-rich end of the lanthanide-magnesium phase diagram is shown in Fig. 1. EXPERIMENTAL PROCEDURES Preparation of Alloys. The lanthanide metals used in the preparation of alloys for this investigation were prepared from the corresponding fluoride by metallothermic reduction techniques previously described by Daane.4 The chemical analyses of the metals used are listed in Table I.

Jan 1, 1965

By G. Thomas, O. Johari

MEIERAN and Richman' have proposed a method for the determination of electron-diffraction patterns from twinned bcc crystals. It is the purpose of this note to present a general method capable of predicting the positions of all reflections from both matrix and twin regions for all possible orientations of cubic crystals. The method has been verified for twinning in explosively deformed copper, and is found to be a very useful tool when used in conjunction with stereographic projection, and dark-field electron microscopy. Let an isometric crystal be twinned in any (hkl) plane. The twinning can be considered either as a reflection in the twin plane or as a 180-deg rotation about the twin axis. In reciprocal space, the 180-deg rotation about the twin axis would place a given point p, q, Y to a new position P, Q,R, such that both will be equally inclined to the twin axis. This is represented in Fig. 1. If the coordinates of three such pairs of points are determined, the matrix of the transformation can be obtained.' It is also seen that such pairs would satisfy another condition, i.e., the vector [p,q,r] plus the vector [P,Q,R] will yield a multiple of hkl, i.e., the indices of the twin plane, because in isometric systems the axis [hkl] is perpendicular to the plane (hkl). This can be simplified by taking the points along the three axes. If a point h2 + k2 + 12, 0, 0 is taken along the x axis, it is seen that its new indices should be so as to yield a sum 2h2, 2hk, 2hl,i.e., 2h(h,k,1). (Also the points h2 - k2 - 12, 2hk, 2h1, and h2 + k2 + 12, 0, 0 are equidistant from the origin and both points make equal angles with the twin axis, i.e., cos-' h/dh2 + k2 + l2.) Similar operations for the points yield respectively, after the 180-deg rotation about the twin axis. Following the method due to Buerger,' the matrix of the transformation is obtained as follows: Using this matrix the indices of any other point in reciprocal space can be determined. This matrix is much simplified for (111) twinning commonly encountered in fcc metals, since are each equal to —1; hence

Jan 1, 1964

By V. B. Dale

San Manuel has been a modern prodigy from its inception to its present status as a block-caving operation producing 33,000 tons per day. The deposit was discovered, prospected, and developed with modern techniques, and nearly all planning was done from results of an extensive drilling program through an average overburden of 670 feet. The Federal Government had a prominent part in the discovery and primary development of San Manuel. The first 17 holes were drilled by the Bureau of Mines, and discovery of the large, low-grade, disseminated copper deposit is credited to the Bureau. A loan fund of 94 million was granted through the Reconstruction Finance Corporation to the San Manuel Copper Corp. for development of the deposit and erection of concentrating and smelting facilities. Then, while the mine was still being developed, the Defense Materials Procurement Agency contracted to buy a large amount of copper at a pegged price.

Jan 1, 1962

By S. R. Gilbert, B. W. Burdett, L. B. Hales

INTRODUCTION Grinding control is currently the most publicized control subject in the minerals industry. Flotation control and crushing control have not as yet received the widespread attention that grinding control has. The primary reasons for this center around the high capital costs of grinding equipment and the fact that grinding mills are extremely energy inefficient. In a typical copper concentrator the energy requirements to produce refined copper are: (1) [Total Energy Distribution (Kw-hr/yr)% Mining4.93 x 10817.3 Mineral Processing9.63 x 10833.8 Smelting and Refining13.9 x 10848.9] One-third of the total energy used for copper production is for mineral processing. A breakdown of this energy into process steps yields the following: [Total Energy Distribution (Kw-hr/yr)% Crushing and Grinding 7.16 x 108 73.8 Flotation 0.76 x 108 7.8 Materials Handling1.78 x 10818.4] It is apparent that comminution operations consume the largest proportion of the energy for processing and as such become major targets for improvements in efficiency. Unfortunately, size reduction operations are extremely inefficient. Efficiency calculations for grinding range from less than 0.1% to about 5% depending on the evaluation criteria. Because of these facts, grinding control has been and will continue to be the focus of major process control efforts. OBJECTIVES AND BENEFITS All benefits achieved as the result of grinding control are derived from modifying physical parameters which improve the efficiency of the grinding mill. These parameters include:

Jan 1, 1982

By Josh B. Taggart

Some twenty to twenty-five years ago we saw the beginning of mechanization in our coal mines, It has now progressed to the point Chat as instruments of production the pick and shovel are things of the past, They had to give way to better and more efficient methods--so mechanization, followed by and combined with automation, has become common-place in our mines. Our coal could not keep its competitive position without this progress Without it our industry could not look to the future with any degress of confidence. Today we lock back and see cases of regained and salvaged markets. We see better coal products being produced at decreasing costs. We see better utilization of manpower and a stronger, more confident industry, Our hats arc off to those mining men, equipment manufacturers, labor leaders, and others who had the foresight to work toward these ends and the ability to bring them about.

Jan 1, 1962