Search Documents

By R. L. Braun

In situ chemical mining of primary sulfide ore requires distribution of injected oxygen in a large underground flooded region of broken ore. Although the maximum oxygen content of the leach solution will be dictated by the operating temperature and hydrostatic pressure, the actual oxygen content will be limited by the geometric dispersion of the injected oxygen, the rate of dissolution of oxygen, and the rate of convective circulation of oxygenated solution within the flooded bed of ore fragments. These factors are examined by a detailed theoretical analysis supported by laboratory experimental data. Typical results of the study illustrate that adequate dissolution of oxygen and circulation of oxygenated solution can be obtained in a column of 4 million metric tons of broken ore by using three gas injection lines and an oxygen injection to consumption ratio of 1.84.

Jan 1, 1974

By C. G. Anderson

The treatment of materials for the recovery of the precious metals imposes many processing and environmental challenges. Sunshine Mining and Refining Company has developed hydrometallurgical technology for recovery of values from refractory precious metal materials. At Sunshine, the industrial focus of these processes has been for the direct recovery of silver without cyanidation. However, recent work has indicated that these technologies can provide a means for treating complex, refractory gold concentrates. This will be the focus of this paper. The practice of pretreating ores and concentrates to make them more amenable to cyanidation is becoming more common. Several pretreatment options for sulfide concentrates containing gold include roasting, bio-leaching, direct pressure oxidation and nitrogen species catalyzed pressure oxidation. The industrial use of nitrogen species catalyzed pressure oxidation has been limited to the Sunshine refinery where its focus has been for the treatment of silver concentrates. Much of the reluctance, outside of Sunshine, to use this type of technology may be based on the perception that high costs are associated with using nitrogen species (e.g. nitric acid, sodium nitrite) to leach. Also, the use of nitrogen species catalyzed pressure oxidation on an industrial scale is unfamiliar to most metallurgical professionals. A summary of pressure leaching processes which utilize nitrogen species as a catalyst is presented in Tables 1 and 2.

Jan 1, 1996

By E. J. Ostrowski

The history of ironmaking dates back to 4000 BC when an open forge was used to produce a sponge iron from a heap of ore and charcoal. Even at this early date, man recognized the need for a fuel and a reducing agent. As time passed, man developed new methods of producing iron; however, these developments were centered on the mechanical aspects, and as a result, a furnace evolved which is now known as a blast furnace. During this period of evolution,' man struggled with the only fuel he knew - charcoal. Some 5600 years passed before any attempt was made to use coal as a fuel for the blast furnace. The use of coal had its beginning in 1619 when Dud Dudley of England first attempted its use in the blast furnace; as a-fuel and as a reducing agent. Coal charged to the top of the blast furnace was not satisfactory. The volatile matter of the coal choked the flues, and also the poor, strength qualities would not support the weight of the ore. Breaking up of the coal compacted the stock column to an impervious mass. One hundred and sixteen. years later in 1735, Abraham Darby of England introduced coke as the fuel and reducing agent in the blast furnace. The acceptance of coke as a fuel for the blast furnace was not spontaneous. As late as 1870, one-half of the pig iron product of the United States was made with anthracite coal, 30 per cent with raw bituminous coal and coke, and 20 per cent with charcoal. However, by 1910, 96.2 per cent of the blast furnace product was made with coke, 2.4 per cent with anthracite and coke admixture, and 1.4 per cent with charcoal. Coke became the standard fuel in the blast furnace. As the art of ironmaking progressed, more pig iron was produced and more coke was consumed. Coke oven capacity was progressively increased. In 1910, there existed approximately 4,000 coke ovens; in 1960 there was approximately 15,000 ovens. This increase in coke oven capacity was necessary to feed the mammoth appetite of modern blast furnaces.

Jan 1, 1962

By H. Lyn Bourne

This is the sixth annual tabulation that gives mineral royalty information. This tabulation, like the previous editions, includes all of the earlier data plus more than 100 new entries. In addition to the royalty information, some entries show the cost of purchased reserves. An article in the December 1984 Kentucky Coal Journal could have far reaching applications. It explains that a banking firm examined the "reserve risk" as one of six factors for analysis when financing a coal project. The firm concluded that coal reserves in Kentucky are worth an average of $0.29/t ($0.26 per st) in-place. The banking firm based this in-place value on its examination of more than 50 leases. In-place value is especially important when trying to establish a mineral value for a given deposit. Ernie Lehman wrote to this author after last year's column appeared and he quoted the lease rates for lead and zinc on federal land. He states, from first hand information, that the leases are based on a percentage of net smelter return as follows: 4% for the first five years, 4.5% for the next five years, and 5% thereafter. The figures used in last year's column ($ per st) came from dividing reported royalty income by reported production. The information reported last year merely translates the results of the net smelter return (as written in the lease) to royalty income per ton to the US Government for the various years. Statistics from the Minerals Management Service of the Federal Government show that oil and gas account for 40% and 53%, respectively, of the Federal royalty income. Coal accounts for about 4% and all other minerals, the remaining 3%. Oil and gas leases are mostly at 12.5%. Most of the coal leases are at 12.5% for surface mining operations and 8% for underground mining. Many of the other commodities are leased at 5%, which has been true for decades. Relative to Federal leases, this year's table shows some percentages and some costs per ton. This was done to give information about the leases and also to show income produced by various commodities. Note that some commodities have a variation in $/ton even though the royalty percent is the same. These variations result from a combination of factors that influence the value of a mineral commodity. The same caveats that have appeared in previous columns relative to factors that influence the value of reserves are restated. The data in the table serve as a general reference and are not intended to present an absolute scale of reserves costs. Several factors influence the value of any mineral commodity. The factors include: • location/transportation (especially significant for most industrial minerals); • market conditions both for the commodity and the company acquiring the right to mine it; *quality or grade of the deposit as it reflects the amount of processing/beneficiation necessary to produce the finished product; and • legislative restrictions and conditions for mining a given material at a given site. These and other factors will significantly influence the cost of reserves. The table shows the range of royalties and the variations in the basis of computing the fees. In a few instances, it gives the cost of acquired reserves based on purchase price and quantitative estimates of the material in place. The table lists the commodity in column 1, and the subsequent columns give the location - either by state, US geographic area or Canadian province; the cost - R for royalty and P for purchased cost; and the year in which the lease or purchase agreement occurred. The last column offers either comments or gives a footnote number for more information. The cost column requires more explanation. The units may refer to cubic yards (/yd), cubic feet (/cu ft), tons (/T, 2000 lbs) or acres (/A). 1 cu yd = 0.7646 m3 1 cu ft = 0.2832 m3 1st=0.9078t 1 acre = 0.4047 hm or m2 In several cases, the royalty is a percentage of the sales price of the finished product and is levied against either the amount produced or the amount sold. We thank those who contributed information for the table: G. Anderson, B. Brown, R. Ganis, K. Hauser, E. Lehman, and S. Sims. Their help is appreciated as are the contributions by those who asked to remain anonymous. This column requires additional information each year. That data must come from people in the minerals industry who have royalty information. What appears in the table is only a very small fraction of the cost information relative to the various commodities and the geographic areas for which there is royalty information. If you find this column of interest and possess information write the author at the address given in the author's box or call 313-462-0005.

Jan 7, 1988

By Donald D. Carr

Four counties in Indiana--LaPorte, Vigo, Vanderburgh, and Switzerland--have potential for deep underground limestone mining because they are in or are near high population areas and have no surface-minable limestone de- posits. The principal markets are Gary-South Bend, Terre Haute, Evansville, and Cincinnati. Minable seams suitable for aggregate are 120 to 150 meters (400 to 500 feet) below the surface in LaPorte and Vigo Counties, about 90 meters (300 feet) below the surface in Switzerland County, and about 600 meters (2,000 feet) below the surface in Vanderburgh County. Some deposits are also suitable for high-purity uses or for cement raw materials, which enhances their potential for exploitation.

Jan 1, 1980

By K. R. Griffin, K. D. Luxbacher, E. C. Jong

The schlieren technique describes an optical analysis method designed to enhance light distortions caused by air movement. The ability to visualize air flow has significant implications for analyzing underground mine ventilation systems, including the enhancement of surveys and models. Currently, traditional schlieren methods are impractical underground. Background-oriented schlieren (BOS) provides a solution to this problem. BOS requires two primary components, a professional quality digital camera and a schlieren background. A schlieren background is composed of a varying contrast repetitive pattern, such as black and white stripes. This background allows the camera’s sensor to capture minor light diffractions, such as those caused by air flow. Although BOS shows great potential for ventilation analysis, sensitivity is an issue. This paper seeks to validate the BOS technique for imaging flow through a regulator by comparing BOS images with traditional schlieren images. This paper also outlines initial experimental efforts used for the validation.

Jan 1, 2011

By E. L. Smith

The development of larger size flotation machines, which started in the 1950's and 1960's, is still proceeding. The largest mechanical flotation machines avail-able today have unit volumes in the range of 38.2 to 43.9 M3 (1,350 to 1,550 cft). Designs for mechanical flotation cells of unit volumes as large as 70.8 m3 (2500 cft) are available. The development of these "giant" flotation machines has been and continues to be motivated by the increasingly critical forces of higher energy and capital costs, decreasing ore grades, and lower (real) prices for the products. As an expansion of a previous publication, this paper reports the further development of an improved mechanism for these large flotation machines. The starting point in the develop¬ment of a new mechanism was the identification of design and performance features of existing machines where improvements were desirable. The result of this work was a mechanism which s is an improvement over existing designs. This new design is characterized by an efficient overhung stator and a vortex profile rotor. The overhung design of the stator removes unnecessary and undesirable baffling in the lowest parts of the tank, while maintaining a quiescent froth layer. This results in lower power consumption and better pulp circulation and solids suspension. The vortex profile rotor acts as an efficient pump, with higher pulp recirculation rates as evidenced by the increased air dispersion capabilities, good solids suspension, and low power consumption.

Jan 1, 1983

By Harrison R. Cooper

The expanding use of digital control in the metallurgical industry is a significant recent development. From 1970 to 1975, approximately forty metallurgical plants throughout the world have adopted automation techniques based on digital controls (see Table 1). Several recent technical reports have described development of digital control systems in concentrator grinding control, 1,2 flotation,3 and smelter operations.4,5

Jan 1, 1975

By N. Edmonson

Introduction An examination of statistics for the world phosphate fertilizer industry since 1980 reveals that there is an increasing tendency for fertilizer-manufacturing activity to concentrate in the areas of phosphate-rock mining. The economic justification for this resource-oriented strategy is clear. The choice between shipping phosphate rock, a comparatively low-value product, and shipping a more valuable finished fertilizer product will always, as long as cost considerations control the decision, be resolved in favor of the latter. The historical existence of substantial phosphate fertilizer-manufacturing capacity, which is oriented toward fertilizer-consuming activities and dependent on imported phosphate rock, shows that narrow cost considerations have not always determined the industry's location. One is entitled to ask what conditions have changed to precipitate the observed relocations, and how permanent are these changes. These questions are the subject of this investigation. Recent trends in world phosphate trade Since 1980, the changing pattern of world trade in phosphate materials reflects an increased interest in manufacturing phosphate fertilizer in rock-producing areas. This is shown by the decline in the total tonnage of phosphate rock being traded worldwide compared to world consumption of phosphate fertilizers (Fig. 1). This relative decline has continued since 1984, which is a longer period of time than could be explained by temporary cyclical market factors. Other evidence corroborates this conclusion. The world's phosphate-rock production is relatively concentrated geographically. In 1992, the top ten producing countries accounted for 92% of world production. Figure 2 shows that the share of world fertilizer production in the top ten rock producers has increased. Yet, further evidence of the direction of trade changes can be had by looking at the pattern of the world's sulfur trade. The first step in the manufacture of any of the widely used phosphate fertilizer products is the acidulation of phosphate rock with sulfuric acid. In most instances, phosphate fertilizer producers have found it economical to produce their own sulfuric acid by burning elemental sulfur, and such operations account for over 60% of all sulfur consumption in the world. 1 Therefore, a concentration of phosphate fertilizer production in rock-producing countries should produce a similar concentration of sulfur imports in the same countries. Figure 3 suggests that there has been a tendency since 1989 for sulfur exports to go increasingly to the rockproducing countries. World phosphate-location strategy is not limited to the alternatives of shipping rock and shipping finished fertilizers. A third alternative consists of producing phosphoric acid, the principal phosphate fertilizer intermediate, at the rock deposit, and then exporting it to market-oriented ammonium phosphate plants. This arrangement offers the importing country the advantages of shipping a more concentrated product while retaining the economic multiplier benefits of a domestic high-analysis fertilizer industry. India, and several other countries, imports merchant-grade phosphoric acid to use as input for the domestic value-added fertilizer industries. For various reasons, the trade in phosphoric acid has never quite lived up to the promise suggested by the foregoing discussion. Figure 4 shows no growth in this trade in the last ten years, either in absolute terms or in relation to the world's consumption of phosphoric acid. Most of the decline in the trade in phosphorous in rock form has been replaced with trade in finished fertilizers. The trade in superphosphoric acid was a growth activity in the late 1970s with the growth in US exports to the USSR, and this trade survived the Carter-era embargoes of US-Soviet trade. However, no additional sizable markets for superphosphoric acid developed, and other suppliers sought out the Soviet market. The problem with superphosphoric acid is that the cost of its incremental concentration is as great as the savings from shipping the more concentrated product. The Soviet market was a special case. The political breakup of the Soviet Union disrupted the superphosphoric acid trade after 1991 and led to a shutdown of the US export capacity.

Jan 1, 1996

By W. M. Luff

The Brunswick No. 12 Deposit is a large polymetallic massive sulphide deposit, occurring in Ordovician age rocks of the Tetagouche Group. Silver is contained within the sulphide zones and economically it is the second most important commodity produced. Brunswick is one of the largest silver producers in Canada. The distribution of silver within the mine is presented, with examples from various levels. The concentration of silver can be shown to be related to both stratigraphy and structure. The mineralogy of silver in mill products is also presented.

Jan 1, 1986

By Harry E. LeGrand

The purpose of this paper is to outline briefly some basic principles of ground-water hydrology that have application in mine hydrology and to suggest that an understanding of these principles can help to forestall problems that are internal as regards mining operations and external as regards damage or harm to surrounding areas. The development of a mine below the water table pleads for accurate values to answers of hydrologic problems. Such internal problems involve both an inflow of water without adequate drainage facilities and a deficiency of water for the required mining or milling use. Methods which give acceptable quantitative values for the solution of many ground-water problems depend on rather simple geologic structures, characterized chiefly by homogeneous rocks of considerable areal extent. Unfortunately, mining operations are likely to be centered where rocks are not homogeneous. Consequently, methods that attempt to show values of the storage and transmitting capacity of rocks in mining areas must be applied with caution. In spite of complex geology and apparent uniqueness of individual mines and quarries, adherence to some concepts of ground-water occurrence and movement should be helpful to mine operators. Mining operations that extend below the water table disturb the natural hydrologic regimen. Prior to development, the water table slopes away from the site of a mine in at least one direction, but as soon as development begins and water is removed from the mine, the mine becomes the apex of a depressed cone in the water table, causing a convergence in the flow of water toward the mine, The behavior of the cone of depression, as it tends to deepen and to increase in areal extent when the mine is deepened, is related to the rate of withdrawal of water necessary to keep the mine dry. The significance and nature of the cone of depression in the vicinity of pumped wells has been studied by ground-water hydrologists for years, but the cone caused by pumping water from mines has generally been less systematically appraised. The geologic and hydrologic settings of mines range greatly so that a simple standard procedure for dealing with problems

Jan 1, 1972

By Robert M. Jimeson

The competitive market potential for solvent refined coal is estimated. Market advantages of solvent refined coal are enumerated. Markets are possible in combustion, railroad locomotion, and carbon electrodes. The combustion characteristics are consistent with expected air pollution restrictions on the emission of sulfur and particulate matter. A market potential of sufficient size exists to support the initiation of solvent refining in the United States. The market analysis should help fuel producers determine the advisability of their undertaking the solvent refining of coal.

Jan 1, 1969

By Frank Florez

INTRODUCTION The Magma mine, located at Superior, AZ, some 100 km (60 miles) east of Phoenix, was discovered in 1875, and was originally mined for silver. The property was acquired by W. B. Thompson who formed the Magma Copper Co. in 1910. Prior to 1965, square-set mining and cut-and-fill methods were used to mine steeply dipping veins at Magma. Ore was moved to the chute by gravity in the square-set stopes and by slusher in the cut-and-fill stopes. The stopes were gobbed with waste from the surface and development headings in the mine. The original square-set method of mining was modi¬fied in order to extract replacement ore bodies in limestones that dipped at 0.52 rad (30°) and varied in thick¬ness from 4.6 to 38 m (15 to 125 ft). Mining of these ore bodies was done by starting at the upper lateral ex¬tremities and retreating to the access openings. The timber square sets and cribbing were used as temporary support, the ground being allowed to cave as the mining retreated. It was found, however, that mining in such a manner was very inefficient; safety hazards increased be¬cause of the blocky and fractured characteristics of the ore and the weakness of the hanging wall. GENERAL DESCRIPTION The undercut-and-fill method of mining is now the primary stoping method used at Magma, and no pillars are left for later extraction. Four ore bodies are pres¬ently being mined. These ore bodies (see Fig. 1) dip conformably at 0.52 rad (30°) to the east with the host limestone and shale sediments, their average strike length is about 122 m (400 ft), and their thickness varies from 9 to 38 m (30 to 125 ft). The ore bodies are separated by varying thicknesses of barren rock. Three of them come together at various elevations primarily in the area of the main feeder fault, but there is an apparent horizontal displacement of about 23 m (75 ft) on the main feeder fault which complicates stope layout. The method as originally applied has been modified in order to mine more efficiently ore bodies that are 9 m (30 ft) or more in thickness. The modification was made to mine all the ore at a given elevation before dropping down for the cut below. The advantage of this is that mining is done on a horizontal plane and there¬fore muck removal can be mechanized. Also multiple faces are made available, thereby improving overall min¬ing efficiency. Another major change made was to carry a 6.7-m (22-ft) face when mining the top sill and a 6.4-m (21-ft) face when mining under fill rather than the former standard 3.35-m (11-ft) high face. This change lengthens the mining cycle and cuts mat prepa¬ration costs in half. Stopes are laid out (see Fig. 2) so that maximum mining panels are available and yet one stope crew will not tie up too much of the ore reserve. One stope might consist of as many as 20 panels, and these panels are divided evenly on opposite sides of the extraction cross¬cut. The stope layout is planned so that the maximum tramming distance to the orepass will not exceed 30.5 m (100 ft) on the crosscut and 38 m (125 ft) in the panel. Stope panel distances are sometimes exceeded when min¬ing to the extremity of the ore body. Tramming distance limits are determined by the efficiency of the com

Jan 1, 1982

By W. C. Baver

Even before the energy crisis developed in the early '70's, considerable research world-wide was being directed towards the development of schemes to both reduce the energy requirements and to reduce air pollution emissions associated with the manufacture of glass. In many cases the sought-for solutions would benefit both objectives. If, for example, the glass batch could be consolidated into such compact forms as briquettes or pellets, they could be preheated in a moving-bed heat exchanger with the high-temperature exhaust gases from the glass furnace. Lower furnace temperatures would thus be possible with a corresponding reduction in heat losses and total fuel consumption. Moreover, the lower temperatures would also reduce the amount of nitrogen oxides generated in the furnace and emitted to the atmosphere. Many tests on consolidated glass batch have been run in many countries. In some cases, extended tests on production furnaces have also been included in the programs. For various reasons, however, none of these has proved to be economically attractive. The current, but temporary lull in the rapidly rising cost of energy which has been with us in recent years, has reduced the push for research for energy saving means for melting glass. A few of the current programs will be discussed. Since raw material requirements, primarily physical in nature, are affected by some of these schemes, these will also be considered.

Jan 1, 1986

By Y. Akin

The floatability of a typical boron mineral, colemanite (Ca2B6O11.5H2O), is shown to be hindered by the presence of montmorillonite type clay minerals. The clay is found to act in the form of slime coating. The mechanism of slime coating is investigated using pure colemanite and clay and their mixtures with an anionic reagent in a microflotation cell. Adsorption, flotation, solubility, zeta potential and scanning electron microscope studies reveal that slime coating occurs through electrostatic attraction. It is possible to minimize the effect of slime coating either at high surfactant concentration or at high pH values.

Jan 1, 1997

By K. A. Smith, X. Cheng

Introduction The separation of minerals from complex sulfide ores by flotation requires careful control of the surface chemical behaviors of the minerals constituting the ores. When the minerals in the ore are finely disseminated, extremely fine grinding is needed to liberate the minerals. Fine grinding is not only an energy-intensive operation, but also influences the flotation behaviors of liberated minerals. Steel balls are commonly used for grinding. Galvanic interactions occur not only between sulfide minerals and grinding media but also among the different sulfide minerals themselves. The electrochemical characteristics of chalcopyrite and pyrrhotite when they are independently coupled with steel grinding media and the relevance of these characteristics to the flotation behavior of these minerals have been exten¬sively studied (Cheng and Iwasaki, 1992: Li and Iwasaki, 1992). These studies dealt with reactions in two electrode combinations, simulating mineral-grinding medium and mineral-mineral interactions. In practice, however, it is more common to involve more than one sulfide mineral. In a study of a pyrite-pyrrhotite-mild steel system, Pozzo et al. (1989) provided evidence that a mineral having an intermediate potential could behave differently, depending on the condi¬tion of galvanic contact. This investigation was undertaken to study a chalcopyrite-pyrrhotite-mild steel system to gain better in¬sight into the mechanism of multiple galvanic interactions. A three-electrode system was chosen to simulate a typical copper sulfide ore. The sulfide minerals were represented by chalcopyrite and pyrrhotite. The grinding medium was rep¬resented by mild steel. The nature of galvanic interactions involving these minerals and grinding medium in an oxygen atmosphere was investigated in a three-electrode combina¬tion, and an attempt was made to relate the laboratory flotation results of pyrrhotite-chalcopyrite ores. Possible implications to the differential flotation of complex sulfide ores are discussed. Key words: Electrochemistry, Flotation, Sulfide ores Experimental The pyrrhotite (57.74% Fe, 0.09% Cu, 4.84% Ni and 36.54% S) sample was obtained from Falconbridge Ltd.'s Strathcona mine, Ontario, Canada, and the chalcopyrite (26.00% Cu, 28.57% Fe, 0.05% Ni and 31.52% S) sample came from Transvaal, South Africa. Mild steel (AISI 1020, 0.2% C) was chosen to represent grinding media. A conventional three-electrode cell was used in the elec¬trochemical measurements. The working, counter and refer¬ence electrodes were, respectively, sulfide minerals or mild steel, graphite rod and saturated calomel electrode (SCE, 0.245 V vs. standard hydrogen electrode). The exposed areas of the sulfide minerals and grinding medium electrodes were approximately the same (0.17 to 0.19 cm2). The preparation methods of working electrodes, the reagents used, the experi¬mental procedures and the instrumentation were the same as those described previously (Cheng and Iwasaki, 1992). Rest potentials of pyrrhotite, chalcopyrite and mild steel were measured individually in 0.05-M sodium sulfate solu¬tions. Combination potentials and galvanic currents of the sulfide minerals and grinding medium were measured in short-circuited two- and three-electrode systems. All poten¬tials in this study are quoted in volts with respect to the standard hydrogen electrode (SHE). To simulate the constantly abrading conditions in a ball mill, rest potentials, combination potentials and galvanic currents were also measured by rotating a porcelain ball against chalcopyrite, pyrrhotite and mild steel electrodes in

Jan 1, 2000

By Sam S. Arentz

Storm King Mines began with the concept that the resource potential of thick and steeply pitching seams of the Grand Hogback region of Colorado could be mined profitably. Two large tracts of private land (fee coal) were acquired; approximately 23.8 km2 near Glenwood and 16.2 km2 near Meeker. Figure 1. Although both properties have substantial coal reserves, one has several distinct advantages in developing a competitive mining operation. Table 1 illustrates these advantages. Among the most important considerations from an economical point of view are proximity to railroad, highway, labor force, quality of coal, seam thickness, and source of water. As can be seen, Storm King Mines' property is adjacent to the main line of the D & RGW Railroad, 1-70, and the Colorado River. The cities of New Castle, Glenwood Springs, Silt, Rifle, Carbondale, and Parachute are within short commuting distances and have available skilled labor.

Jan 1, 1983

By James R. Taylor

24.1-BASIC PRINCIPLES JAMES R. TAYLOR The basic principle involved in maintenance is protection of a company's investment through the care and upkeep of machinery and equipment, buildings and properties. Maintenance is justified through making available to the maximum extent possible the facilities, machines and services required by the operating and production departments in their function of insuring optimum return on investment. As the size, cost and complexity of the equipment used in modern mining and related industries increase, operating and production personnel become proportionately more dependent upon maintenance engineering for its contribution to the overall profit picture. A greater part of the total cost now is charged to maintenance and the maintenance group becomes a major company unit. Basic maintenance concepts have changed very little over the years, but the advances in methods and technique have brought about tremendous challenge and have in consequence assumed a greater influence and importance. The older "sledge- hammer" methods have given way to engineered and scientific systems, while the accelerating development of intricate equipment and machinery, and the instrumentation essential for control, demand an ever-increasing upgrading of maintenance personnel. Regardless of this growth in importance, cost and complexity of the maintenance function, it is important to remember that it is not an independent and self-sufficient unit, but is a necessary facet of the entire plant operation and is subject to the company's total production effort. As one part of a team, maintenance is successful only when functioning cooperatively wit11 the other departments it serves Close collaboration in all areas served by n maintenance group is essential if it is to function fully to protect the company's investment, whether the assets are in machinery, materials or people. 24.1.1-MAINTENANCE SCOPE Maintenance functions may be grouped into the two general classifications of primary and secondary. This grouping constitutes a basic division although, in practice, the scope of the maintenance activity varies in each mine or plant and is subject to differences in size and type of operations and company policy, plus the influence of precedent, which may be sectional or industry-wide. The primary functions supply the justification for the maintenance engineering department. while secondary functions are variable according to plant conditions and needs. The primary functions may be categorized as follows: 1. Maintenance of existing machinery and equipment. 2. Maintenance of existing buildings and grounds. 3. Inspection and lubrication of equipment and machinery. 4. Utilities generation and distribution. 5. Alterations to existing equipment and facilities. 6. Installation of new equipment and facilities. 7. Instrumentation.

Jan 1, 1973

By P. Somasundaran

Electrochemical properties of silicate minerals govern their behavior in processes such as flocculation and enhanced oil recovery that can occur at elevated temperatures. Knowledge of these properties as a function of temperature can be helpful in developing an under- standing of the role of these interfacial properties at non-ambient temperatures. Zeta potential of Sodium Kaolinite and quartz has been determined as a function of pH and temperature in this work. Both systems exhibited markedly different behavior at higher temperatures and also exhibited significant hysteresis. Results are examined in terms of possible dissolution of the minerals and surface reactions at different temperatures.

Jan 1, 1986

By S. A. Shoop

Vibrations from thirty-seven blasts were monitored using three seismographs. Data analysis showed that square root, cube root, and site-specific scaling predicted ground motion with equal reliability. Imprecise prediction is caused by geologic and blast variability, particularly inaccurate delay times. Vibrations at some mines were consistently lower or higher than vibrations reported in the U.S. Bureau of Mines studies. Regulations based on USBM data could then be too conservative or lenient. Predominant vibration frequencies were generally less than 50 Hz, but several blasts had significant energy up to 125 Hz. In some cases energy was concentrated at frequencies corresponding to delay intervals.

Jan 1, 1983