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By Rainer Kahrger, Guenter Ewald

The topic of this paper is to demonstrate the development of mobile crushing installations from throughput capacities of 200 tons per hour, to mobile and semi-mobile large- scale crushing installations with throughput capacities of 3,500 up to 10,000 tons per hour. Basis are approximately 140 mobile and semi-mobile crushing installations which have been built since 1956 or were on order in 1987. [FIGURE 1. Total Number of Mobile or Semi-mobile Crushing Plants] Sixteen of these approximate 140 mobile and semi-mobile crushing installations have a throughput higher than 3,500 metric tons per hour. Mobile crushers were developed more than 25 years ago in Europe for the quarry industry, which had special requirements - fairly low capacity, mostly softer and less abrasive material (limestone) and full mobility. So the majority of the units up to about 1980 were fully mobile, having a fairly small capacity and utilizing mainly hammermills, double roll and impact crushers. It was only when the mining industry and large quarries picked up the concept in the early 1980's that a drastic increase in capacity took place. Even though the first mobile crusher in a true mining application went into service in 1971, it took another 10 years before it proceeded in a larger scale. This trend was obviously supported by the drastic fuel price and energy cost increase in the mid 70's. Since the cost of conveying material is substantially cheaper than trucking it, the increase in energy cost, combined with increasing haulage distances provided a strong incentive. Figure 1 shows clearly the throughput distribution of all crushing installations and it also shows the sharp increase in the throughput capacities within the last few years.

Jan 1, 1989

By D. A. Barr

The Galore Creek copper deposits lie in the Stikine River district of northwest British Columbia, a largely undeveloped region which possesses rugged mountain terrain with deeply incised valleys which support dense undergrowth. Timberline is normally at about 1300 m (4265 ft) and the higher levels contain glacial systems and large ice fields with peaks rising to 3500 m (I 1,483 ft). A location map of the Galore Creek area is given in Fig. 1. In the late 1920s, field parties under the direction of F.A. Kerr (1948) of the Geological Survey of Canada completed geological surveys of the district. Primary access for this work was provided by riverboats on the Stikine and Iskut Rivers. Kerr's party had to contend with extremely difficult travel conditions that required considerable backpacking, augmented by pack dogs in certain cases. Absence of ground control led Kerr to establish his own topographic maps. Among the geological data established was the recognition [ ] that small bodies of orthoclase porphyry in the district are often well mineralized with base metal sulfides. Copper mineralization in the upper Galore Creek valley, a tributary of Scud River which flows into the Stikine River some 80 km (50 miles) south of the small settlement of Telegraph Creek, was originally discovered during aerial reconnaissance by M. Morrison and W. Buckholz while prospecting for Hudson Bay Exploration and Development Co. Ltd. in 1955. Malachite- and azurite-stained out- crops in several parts of the Galore Creek watershed are clearly visible from distances of 500 m ( 1640 ft) or more. Claims were staked in 1956 and the copper showings were partly mapped, trenched, and investigated by shallow diamond drilling. Because of the remoteness of the area and higher priorities for expenditures elsewhere, no further work was undertaken and most of the claims were allowed to expire.

Jan 1, 1991

By W. N. Key, L. C. Acton

Freeport Minerals Co.'s Ertsberg East underground operation typifies the cliche that all the "easy" ore deposits have already been found. Despite its location, in a remote Indonesian rain forest, and the effort and expense required to develop the ore body, the mine was developed on schedule, under budget, and with minimal disruption of the adjacent open-pit mining operation. Development of the Ertsberg East copper, gold, and silver ore body took less than two years after exploration drilling and feasibility studies were completed. In addition, the project came in well under its targeted budget of $101.5 million. Initial block caving and underground ore production took place last year. About 3 kt/d of Ertsberg's total 8.6 kt/d ore production is currently mined underground. This will increase to 5.4 kt/d by year's end. Underground operations will eventually account for all of Ertsberg's production, as the open-pit ore body is depleted. Freeport Minerals is currently studying the possibility of increasing total Ertsberg East production tonnage to 10.5 kt/d by 1983. The equipment, work practices, and development methods at Ertsberg East (Gunung Bijih Timur or "GBT"), while typical of other block caving mines, were modified for rapid mine development and maximum ore recovery. Located in the Irian Jaya province, "Gunung Bijih" is the Indonesian expression for the Dutch "Ertsberg" that in English means "ore mountain." "Timur" means "east" and relates to the fact that the new underground ore body lies about 1 km east of the open pit. The area surrounding GBT is mountainous, with peaks from 4.2-5 km. The underground mine adit and surface plant are located at a 3.6-km elevation. The mine, 4° south of the equator, gets 5.1 m/a of rain, but has an average daily temperature of only 13°C because of the mine's elevation.

Jan 7, 1981

By Roy Moreby

Dartbrook Coal is currently developing a modern, high production (3.5 Mtpy) single longwall retreat coal mine in the Upper Hunter Valley of NSW, Australia. The location is a green field site, with no surface or underground mining operations ever having taken place in the initial target seams. The coal seams of the working section, floor and immediate roof average a total 20 m in thickness and contain a mixed carbon dioxide 1 methane seam gas ranging from 6.5 to 11.0 m3/t. The target seams have undergone detailed characterisation to assess hazards and quantify the controls required to minimise associated risks during mining. Risks identified during initial phases of exploration and 14 months of development include: • High gas emission during development. • Probable high longwall gas emission rate. • Medium to high propensity to spontaneous combustion • Possible carbon dioxide outbursts on geological structures • The main controls currently in place include, • Mine geometry and ventilation circuits • Exploratory drilling ahead of development • Intensive in seam pre drainage • Goaf drainage plants • Comprehensive monitoring systems • Underground environmental management plan. • Pre mining predictions of gas emission provided an indication of gas emission rates which have proven to be an under estimate. Current mine issues are; • Both surface fans are at full capacity (250 m3/s 1.4 kPa) • Both gas drainage pumps are at full capacity (2.3 m3/s) • Alternatives to extensive in seam pre drainage being investigated (surface to in seam hydro fracture) Review of mine design underway This paper describes the seam characterisation, assessment of mining hazards and summary of the management plan developed to address the risks identified. A review of current mine underground environmental performance is also provided

Jan 1, 1997

By J. Wagner, P. Eger, G. Melchert

In 1991, LTV Steel Mining decided to close an open-pit taconite mine in northeastern Minnesota using a passive-treatment approach consisting of limiting infiltration into the stockpiles and wetland treatment to remove metals. More than 50 Aft (55 million st) of sulfide-containing waste had been stockpiled adjacent to the mine during its 30 years of operation. Drainage from the stockpiles contained elevated levels of copper, nickel, cobalt and zinc. Nickel is the major trace metal in the drainages. Before the closure, the annual median concentrations ranged from 1.5 to 50 mg/L. Copper, cobalt and zinc are also present but they are generally less than 5% of the nickel values. Median pH levels range from 5 to 7.5, but most of the stockpile drainages have pH levels greater than 6.5. Based on the chemical composition of each stockpile, a cover material was selected. The higher the potential that a stockpile had to produce acid drainage, the lower the permeability of the capping material required Covers ranged from overburden soil removed at the mine to a flexible plastic liner. Predictions of the reduction in infiltration ranged from 40% for the native soil to more than 90% for the plastic liner. Five constructed wetlands have been installed since 1992. They have removed 60% to 90% of the nickel in the drainages. Total capital costs for all the infiltration reduction and wetlands exceeded $6.5 million, but maintenance costs are less than 1 % of those for an active treatment plant. Because mine-drainage problems can continue for more than 100 years, the lower annual operating costs should pay for the construction of the wetland-treatment systems within seven years.

Jan 1, 2000

By Kenneth V. Luza

Underclays are associated with coals found in eastern Oklahoma. These coals are of bituminous rank and are in beds of Middle and Late Pennsylvanian age. They occur in an area of about 8,000 mi2, within the southern part of the Western Interior Coal Region. The rocks of the area consist of a thick series of sandstones, shales, underclays, coals, and limestones. There are more than 40 beds of bituminous coal in the state, but only 12 of these have been mined over large areas. The 12 are, from oldest to youngest: the lower Hartshorne, upper Hartshorne, McAlester (Stigler), upper McAlester, Cavanal, Rowe, Secor, Weir-Pittsburg, Mineral, Croweburg, Iron Post, and Dawson coals. None of the underclay associated with these coal beds is known to have been exploited commercially. Nine samples of underclay from beneath the Croweburg coal were collected over a wide area in the eastern Oklahoma coal field for evaluation. The underclays are usually shades of gray with brown or orange iron-oxide staining. Although soft and plastic in some places, they are generally hard and slickensided, and contain abundant carbonized plant fragments. Bulk chemical analyses, clay mineral identification, and ceramic tests were performed on each underclay sample. The dominant clay minerals are illite and kaolinite with some mixed-layer clays. Silica (Si02) ranges from 45.96% to 59.03% and alumina (Al2O3) varies from 14.98% to 23.60%. The U.S. Bureau of Mines Tuscaloosa Re- search Center conducted the ceramic evaluations. Samples were slow fired from 1,000°C to 1,250°C in 50°C intervals. Bulk density, color, percent linear shrinkage, and percent absorption were determined for each slow fire test. All Croweburg underclay samples are best suited for building brick.

Jan 1, 1999

In west-central England, 183 m (600 ft) below the Cheshire Plain, lie the evaporate remains of an ancient sea. Stratified into nine beds of salt separated by marl bands, the deposit stretches roughly 21 x 12 km (13 x 7.5 miles) and contains reserves of rock salt estimated as high as 400 Gt (440 billion st). The Winsford Salt Works, one of England's largest underground mines, extracts the Cheshire salt and sells it for use as animal feed, for application in de-icing programs, and as a key ingredient in sodium fertilizer production and a variety of other industrial processes. Because road de-icing is an important market, demand for Winsford's salt varies with the severity of British winters. Design capacity is 2.25 Mt/a (2.5 million/year). However, production peaked at 2 Mt (2.2 million st) in 1987. Over the past five years, annual production has averaged 1.6 Mt (1.7 million st). Run-of-mine salt averages 92.5% NaCl, mixed with about 5% marl. The marl was formed by deposition of a reddish dust blown into the Cheshire basin from deserts in what is now central Europe. Association with the marl adds color to the salt, producing tea-colored crystals that resemble unrefined sugar. Adaptation of coal mining systems Rock salt mining in the Winsford district began at least as early as 1682. Mining at the current site dates from 1844 but was shut down from 1892 to 1928. In the early 1900s, a series of catastrophic collapses and floods plagued and ultimately shut down then-active Cheshire brine-salt producers. Subsequently, Winsford reopened and began adapting mechanized underground coal mining systems to rock salt extraction.

Jan 1, 1992

By Keith Kern, Joseph P. Matoney, Kenneth E. Harrison

INTRODUCTION "When there is an economic need for cleaning fine coal to low ash, the technical difficulties will be solved. The utilization of fines below 28 mesh will determine the degree to which beneficiation will be carried. Whenever these "fines" will be processed into smokeless fuels or used as new materials for the chemical industry or for hydrogenation, a higher degree of cleaning will be economically justified”.(l) This quote from the 1943 edition of Coal Preparation correctly anticipates the use of fines as "smokeless" fuel or as a future component of boiler fuel and a1 so as a raw material for the chemical industry or possible new uses such as for internal combustion engines. Fines utilization depends upon an economic need and possibly neglects to say an environmental need or a government funded development program as the prime mover for development. In any event, this conference is faced with the "need" to beneficiate fine coal and undoubtedly among the processes that will be discussed are the unit processes of future coal preparation facilities. HISTORY OF DEVELOPMENT It is interesting to look at where we have been, where we are now and where we possibly will be in the future. In order to put some limitations on this presentation, we will for the most part confine our remarks to this century and particularly developments since World War II. We also will limit our remarks to the preparation of bituminous coals in the United States, and for the most part, to wet processes. In the 1943 edition of Coal Preparation, clues to the processes being used to beneficiate coal are contained in the Chapter Titles. An extensive treatment of "Hand Picking" was contained in Chapter 8. Chapter 7 contained a description of methods for "Preparation at the Face". Chapter 18 dealt with "Slurries, Sludges, Slimes and Water Treatment" and defined this size of coal as being the size at which "Jigs and Chance Cones are usually dispensed with and other methods such as the hydrator, tables, and froth flotation are employed. Generally the limit is 1.7mm to 3.4mm. In the days before World War I1 the then modern bituminous coal preparation plant contained jigs or Chance Cones for the coarse material +6.35mm, concentrating tables, launders or some classifier type of coal washer for the fines (6.35 x 0.15mm), and froth flotation as the only method of cleaning minus 0.15mm coal which was used very sparingly.

Jan 1, 1988

By S. R. S. Sastri, G. V. Rao

Because the nickel contained in lateritic nickel ores does not occur in an identifiable mineral, it is generally accepted that these ores are not amenable to beneficiation. For this reason, such ores are either directly smelted or leached to extract the nickel. In this paper, a process for enriching the nickel content of the lateritic nickel-bearing chromite overburden, normally a waste material generated during chromite mining, is described. In this flotation process, a nickel preconcentrate containing >1.1% Ni was obtained from material containing as little as 0.36% nickel. An empirical relationship was developed to predict the grade of nickel that can be obtained.

Jan 1, 1997

By Andrzej Frycz

The development of the theory of reliability creates more and more possibilities of its application in different domains of technics. The methods evolved to date, both theoretical analysis of the engineering systems and investigations techniques in practice can be put in use a d translated from one science to another one [1,3,7J . It needs to consider the specific of the different sciences. But it is necessary to take account of the specific characteristic of each line of science. The reliability investigations needs, justified in technical and economical viewpoint appear too in mine ventilation technics and systems. The mine ventilation system is a complex of workings and ventilation plants, which serve to supply the networks with a sufficient quantity of fresh air from the atmosphere and then carry off the vitiated air. The condition necessary for a ventilation system to perform definite task is the correctness of all its elements, that is a capability to preserve in a expected range of tolerance, the values of basic parameters. The element of a system which ceases to work correctly is called failed. The basic parameter of all elements of a ventilation system is the characteristic curve, pressure loss - air quantity relation ?p = f(Q) (1) This curve concerning the workings and ventilation stoppings is given by the formula: ?p = R Q2 (2) thus the resistance R (of the workings or stoppings) may be taken as a characteristic parameter. The course of this characteristic f(Q) will be given in a complex form in concern with fans, ejectors, air curtains. The failure of a ventilation plant or working is therefore an event, relying on a change of its characteristic, thus for a definite time the losses of the plant's or working's capability set in for the fulfilment of its function in the ventilation system. The failures are divided in viewpoint of their nature on sudden and progressive. We are in concern with the sudden failures which are characterised by a violent change of the ventilation element parameter for example a fall of roof, or silt up of the working, opening or failure of the ventilation stopping, the stand-off of the fan or ejector etc. We distinguish two kinds of failures on the viewpoint of their action on the air flow: the "opening" failure and "closure" failure. The prevalence of factors, which causes the failures is in connection with the intersection in the mine of roads of haulage, roads of underground transport of materials and lining elements, roads of people motion with airways, as also depends on the rock pressure reaction on the workings lining and plants placed there. Other factors will yet be decisive upon supply decay of fans, ejectors and curtains. The appearance of failure of the ventilation element is therefore of random nature and determines the change of the element, when it is passing from operating state to unoperate. Qualifying the ventilation system element to one of these states is easy, when the characteristic change is clear and occurs suddenly, that was assumed. The investigations conducted by Patrushev and Karnaukh [91 in concern with the ventilation stoppings resistance change in the case of their opening, had shown the possibility of resistance value change in the range of several hundreds higher.

Jan 1, 1980

By Robert A. Jankowski, Charles Babbit, Natesa I. Jayaraman

The Bureau of Mines, together with Foster Miller Inc., has developed a compact, high pressure scrubber system for continuous miners. Powered by water energy alone, it can move up to 1.4 m3/s (3,000 cfm) of air and scrub a significant amount of the dust in the air. Full- scale model and underground trials showed excellent performance. Operated at 8.3 x 10[3] kPa (1,200 psi) and 38 L/m (10 gpm), the scrubber has reduced operator and return dust concentrations by more than 50 pct. This paper describes the full-scale model development, system fabrication and underground evaluations of this novel and inexpensive scrubber.

Jan 1, 1989

By Hidehiko Nakazaki, Sakuro Murayama, Mitsuo Miura, Masaaki Endo, Yoshiaki Hurukawa

This is a detailed report of construction experience of a tunnel built by a shield of rectangular cross section with reinforced concrete segments. In case of tunneling by a rectangular shield, there were the following problems to be overcom by technical developments; (1) structural joints between segments, (2) complexity in erector handling, (3) exact steering operation of shield and perfect filling of tail void left around the lining erected. Such rectangular shape of cross section has various advantages comparing the conventional circular shape as follows; (1) effective utilization of excavated space, (2) greater overburden depth, (3) stability of lining against unbalanced earth pressure caused by neighbouring construction.

Jan 1, 1979

By R. N. Tempel, M. Lengke

Arsenic sulfide (orpiment, realgar, amorphous AsS and As2S3) oxidation rates were measured in mixed flow reactors at pH-2 and at temperatures of 25 ° to 30 °C. The measured oxidation rates for arsenic sulfide solids by dissolved oxygen are in the range of 10-11 to 10-10 mole m-2s-1 after being normalized to final surface area. The order of arsenic sulfide oxidation rates decreases at pH-2 from amorphous AsS > realgar > orpiment > amorphous As2S3. Compared to pyrite, arsenic sulfide oxidation rates are approximately 1.5 to 15 times slower than pyrite oxidation rates at pH-2.

Jan 1, 2003

By Qiu Rongqing, Huang Kongxuan

This paper summarizes the results of treating arsenic­bearing refractory gold concentrates from six gold Mines of Chi­na by using bioleaching method and discusses the conditions of bioleaching processes. Results indicate that Thiobacillus fer­rooxidans T-3 Strain can intensely oxidize arsenopyrite, how­ever oxidation rate and degree differe considerably for different. During initial inoculation of 5 X 106 cell/ml. FeH contration generaly 4 -8 g/1. Sample size > 80 % - 3 20 mesh and pulp density 10 % ( w /v) and leaching time ,1 -- G clays, arsenic re­moval rate reached more than 80 % , Fe extraction ,10 - GO% desulfurization rate 30-45% after bacterial oxidation of 4-G days.

Jan 1, 1992

By R. W. Carn, D. Vincent

During the next 15 years, US coal production is expected to double, with the increased production evenly divided between the East and the West. Along with greater production, coal export markets should increase dramatically from East, West, and Gulf Coast ports. The annual overseas export capacity of US coal-loading terminals is expected to rise from 147.1 Mt (162.1 million st) in 1981 to a minimum of 278.1 Mt (306.6 million st) in 1985, according to the US Maritime Administration. Increased coal production and use will lead to more development of import and export terminals, a vital link in the coal transportation chain. With continually escalating capital costs and the competitive markets that the terminals will serve, a well designed and efficient terminal is necessary. This article begins a two-part series that presents concepts presently used in coal export terminal design. Part I looks at site selection factors and equipment needs, while Part II will examine environmental considerations in building a terminal as well as typical capital and operating costs. The world is nearing the end of the oil era. In a few years oil will not be available to sustain the growth rate and increasing standard of living we have known in our lifetime. The big question is what energy era are we moving into? With the decline of readily available oil reserves and rapidly increasing prices, many countries are trying to switch to alternate energy forms. While intensive efforts to find new oil reserves continue, alternate energy sources such as natural gas, coal, synthetic fuels, nuclear, hydroelectric, solar, and wind power are being developed. Recent indications are that coal is expected to bridge the energy gap over the next 25-30 years until the technology and economics of the alternate energy forms reach satisfactory levels. Use of coal for energy is receiving strong attention due to its long-term availability (200-300 years minimum), relative ease of development, and its low cost per unit of power produced. By the year 2000, it is expected that 25% of world energy supply will be met by direct coal combustion and possibly another 5-10% by synthetic fuel from coal. Coal's expanding share in the world energy market, along with an increase in coking coal requirements, will result in a large increase in the world's seaborne coal trade. Recent statistics and projections for the future are shown in Table 1. This phenomenal development rate includes increases in both coking and thermal coal requirements. Because of the rapid increase of seaborne coal trade during the last 10 years and the even greater projected increase of trade to 2000, various sectors of the coal industry are faced with enormous technical challenges and huge investments in equipment, land, transportation systems, and port facilities. Very large bulk terminals are under development throughout the world. Latest surveys indicate that there are about 30 new coal export and import terminals under consideration and at least 30 existing terminals have expansion programs planned or underway. With the high cost of borrowed capital and rapid inflation rates there is great emphasis on new planning and design techniques to minimize capital and operating costs of coal transportation systems. Terminals A total coal supply system can be considered to consist of one or more mines; a train, barge, truck, or other haulage system; an export terminal; a fleet of bulk carriers; a receiving terminal; and possibly, local inland distribution networks that include barges and railways. Terminals, though only a small link in the total transportation system, play a key role in overall system efficiency. At ports or inland distribution centers, terminals act as transportation links bringing trains, ships, barges, or trucks together for cargo transfer and temporary storage. A well-designed terminal can provide maximum independence between two modes of transportation and optimum freedom for intermodal interference. A terminal acts as a buffer between the two transportation modes by providing sufficient storage capacity so a ship need not wait for its cargo on, for example, a train-by-train basis, but can load immediately from the ready stock. Similarly, a train need not wait for a ship to unload its contents but can dump immediately into storage. A terminal also can be used to properly mix various types of coal to satisfy a buyer's requirements. Consider the relative value of various production and transport segments for a typical steam coal

Jan 6, 1983

By R. A. Metz, Willard C. Lacy

Rather than attempting to present a summary of the many and highly varied papers that have been presented at this symposium on sampling and grade control, I will attempt to extract the general philosophy of analysis and approach, and attempt to identify the trend of future developments. First, the term "sampling" is used with its broadest connotations. A sample consists of a representative portion of a larger mass, and must represent the mass not only in the grade of contained metals or minerals, but also in all other respects in terms of mineralogy and mineral quality (1, 5), deleterious materials, recoverability of economic components, physical behavior, geophysical response (1), and even archaeological and environmental aspects (7, 11). The sample must be taken from a locality and in such a manner and quantity that it is representative of the larger rock mass. This calls for complete and accurate geological control and an understanding of the nature and distribution of the contained chemical and physical elements and a record of the effectiveness of the different sampling methods. Second, value of a given mass of ore material is based upon its profitability - the difference between recoverable value and costs to achieve recovery, beneficiation and sale. There is a strong movement in mining geology control toward more complete analysis in determining cutoff grades and in grade control, as illustrated by the kriging of metallurgical recovery factors as well as grade at the Mercur Mine (8). To achieve a "profitability factor" as a guide for economic mining practice requires further integration of: 1) the value of contained metal or mineral, 2) percentage recovery of values, 3) dilution of ore with waste rock, 4) addition to, or loss of value as a consequence of by-product materials or deleterious components, 5) cost of producing a saleable product plus minimum profit to justify the effort (cutoff), and 6) cost of land restoration (7, 11). All these parameters vary with the rock type, rock structure, mineralogy, depth, geometry, mining and metallurgical methods, but they must be sampled and analyzed if sampling and grade control are to reflect profitability. A wide variety of deposits has been presented at this symposium; each deposit with its own problems and special solutions. Deposits containing high unit-value components, e.g. precious metals and diamonds, present special problems in the obtaining of accurate samples and generally require statistical analysis control methods or may disregard or modify occasional high or occasional low values, based upon experience (12). Grade control may be accurate for the long term but may vary for the short term. Bulk sampling is always essential. Deposits containing metals or minerals with low unit value are very sensitive to transport costs, and they are often very sensitive to small amounts of deleterious components or differences in physical or chemical behavior. Problems of sampling and grade control change with the genetic type of deposit, with the stage of deposit development and with the size of the information base. Precious metal epithermal deposits (2, 6, 8), because of rapid vertical zonation and erratic lateral distribution of values, have always been difficult to evaluate and maintain grade control and ore reserves. On the other hand, evaluation and grade control are relatively easy in bulk-lowgrade deposits (4, 13). However, these deposits generally have a low margin of profit and are sensitive to mining and beneficiaton costs, price fluctuations and political costs. Industrial mineral deposits (5) often must be evaluated on the basis of their behavior, rather than by chemical analysis. Environmental impact generally increases with the scale of the operation, but certain elements or minerals have especially high impact effects (7, 11). In the exploration phase there is no production control of sampling procedures and careful geological observations are particularly essential. The greatest number of problems is related to the oxidized outcrop where the chemical environment of the ore body has changed and the contained values may have been enriched, depleted or values left unchanged (2, 6). Present evidence suggests that gold values may be very mobile under certain conditions (2, 6) and stable under others. Everything must be sampled in detail. Principal values and by-product or deleterious elements may vary dependent upon their position within the soil profile. Such factors as geomorphic position, erosion rate, vegetation, climate, etc., may affect the interpretation (1, 3). During the development phase it is equally easy to overtest, to have "paralysis by analysis," as to undertest (3, 6). Bulk samplng and testing are

Jan 1, 1992

By Gordon T. Austin

Garnet is the general name for a family of complex silicate minerals having similar physical properties and crystallizing in the isometric (cubic) system. All garnets have the same general chemical formula but vary greatly in chemical composition. The name garnet is derived from the Latin word granatus, meaning like a grain, which refers to the mode of occurrence wherein crystals resemble grains or seeds embedded in the matrix. GEOLOGY Mineralogy Chemical Properties: The general formula for garnet is A3B2(SiO4)3, whereas A can be calcium, magnesium, ferrous iron, or manganese, and B can be aluminum. ferric iron, or chromium, or rarely titanium. The formulas and names of common garnet species are: [ ] Almandite and almandite-pyrope solid solution garnets are the best abrasive types, but andradite, grossularite, and pyrope also are used. All species of garnet have been used as gemstones. The structure of garnet consists essentially of isolated SiO, tetrahedra connected by oxygen-cation-oxygen bonds through the distinct A and B group cation sites. Within this structure magnesium, ferrous iron, and manganese easily interchange and substitute for each other in the A cation position, and calcium does so less readily. Additionally, aluminum, femc iron, and chromium substitute for each other to a limited extent in the B cation position. This ability to substitute or exchange ions without changing the crystal structure is called isomorphism, and garnet is one of the finest examples of a isomorphous series. Because of this isomorphism there is complete solid solution between certain garnet species but not between others. Fig. 1 illustrates these solid solution relationships. Some rare species of garnet are known that illustrate the wide range of cation substitution that the garnet crystal structure can accommodate. They include: [ ] These rare species are not of interest for industrial applications, but can be of interest to mineralogists and the gem industry. Physical Properties: Garnet displays the greatest variety of color of any industrial mineral. Garnets have been found in all colors except blue. For example, grossularite can be colorless, white, gray, yellow, yellowish green, various shades of green, brown, pink, reddish, or black. Andradite garnet can be yellow- green, green, greenish brown, orangy yellow, brown, grayish black or black. Pyrope is commonly purplish red, pinkish red, orangy red, crimson, or dark red; and almandite is deep red, brownish red, brownish black or violet-red. Spessartite garnet can be red, reddish orange, orange, yellow-brown, reddish brown, or blackish brown. A few garnets exhibit a color-change phenomenon. They are one color when viewed in natural light and another color when viewed in incandescent light. Because of the great variation in color of garnet within each species and similarities in color between garnets of different species it is recommended that garnet identification not be based on color alone. The Mohs hardness of garnet varies from 6.5 to 9.0. Grossularite and uvarovite have a hardness of 6.5 to 7.5; andradite is 6.5 to 7.0; and pyrope, almandite, and spessartite are 7.0 to 7.5 in hardness. There are reports of almandite having a hardness of between 8.0 and 9.0. As with hardness, the specific gravity of garnet varies considerably. The specific gravity may be as low as 3.2 or as high as 4.3 depending on chemical composition. Garnet crystallizes in the isometric system with rhombic dodecahedra and trapezohedra the most common forms. Crystals also form in combinations of dodecahedra and trapezohedra or either of these in combination with hexoctahedra. Crystals can have cubic or octahedral faces, but these are rare. Under favorable conditions of formation garnet will crystalize in nearly perfect forms that rival study models. Garnet also forms as irregular blebs, grains, knots, or masses, with or without distinguishable crystal faces, and as coarse- or fine-grained granular masses that appear to be totally lacking in crystal form. The fracture of garnet also shows great variation. In some garnets, particularly those that are well crystallized and glassy in appearance, the fracture is subconchoidal to conchoidal. Other, more poorly crystallized garnets exhibit a fracture that can only be described as uneven. Garnet occasionally has an indistinct dodecahedral cleavage. Certain species of garnet from specific locations have a pronounced laminated structure consisting of planes of weakness along which parting takes place. This parting may resemble cleavage, but since it is mechanical and not related to crystal structure it is not true cleavage. The optical properties of garnet are sensitive to even small changes in chemical composition or strain in the crystal structure. For this reason each species does not have a single index of re- fraction (IR) but has a range of 1Rs. Uvarovite in its pure form should have an IR of 1.870, but in nature samples vary from 1.74

Jan 1, 1994

By Richard R. Klimpel

Part 2 of this article continues the discussion of a 10-year, multimillion dollar research and plant testing program on slurry rheology and grinding circuits. The first part of the article (ME, Dec. 1982) presented the basic concepts identified by the research and some laboratory test results. This section Illustrates typical Industrial scale test results and Identifies some industrial operating implications of controlling rheology by different methods. At least four controllable factors decide the rheological character of a slurry-slurry density or percent solids, particle size distribution, chemical environment, and slurry temperature. The second factor has two interrelated facets: the shape of the particle size distribution which controls packing behavior of the solids, and the fineness of the distribution. Finer particles increase interparticle forces and viscosity. As indicated in Part 1, during a given grinding test it is possible for all four factors to change. However, regardless of the particular settings of the four factors in a given test, if the resulting rheological character is either dilatant, pseudoplastic, or pseudoplastic with yield, the associated breakage rate is correlated with the current rheological character. It is obvious, for example, in batch grinding tests run at constant percent solids, that the second and third factors, where appropriate, are changing during grinding because size distributions are changing and the production of fresh surface area takes up unadsorbed chemical. Thus the corresponding rheological character change in batch tests with increasing grind time would be dilatant to pseudoplastic to pseudoplastic with yield. The degree to which this transformation occurs depends on the changing setting of the four factors over grinding time. In continuous grinding tests, one might logically expect that dramatic changes in any one of the four factors will be less likely to occur. It will be shown that continuous mill operations offer some unique opportunities to take advantage of possible rheological transformation by more direct operational control of the settings for the four factors. One extra observation noted in the rheological studies was the variability in the location and extent of region B (pseudoplastic behavior) for the various coals and ores tested. The location of region B was usually in the region of 45-55% solids by volume and was of the extent of 0-8%, or 2-11% with chemical addition. The corresponding increase in net production ranged from 0-10% in region B and 0-21% in region B&apos;. When region B is small or zero (no pseudoplastic character is exhibited) no increase in production will be observed and the use of chemicals is often marginal. There are several reasons for some materials exhibiting this quick transformation from dilatancy to high yield values often at surprisingly low percent solids by volume such as 30%. One condition identified was for materials containing high levels of viscosity-producing elements such as- carbonates or clays. A second condition documented was for materials that exhibited unusually fine primary fragment distributions (fine B1,; curves). This corresponds to materials having small y values of < 0.5. Such a slurry developed a yield value quickly during grinding because of the rapid buildup of fines. A related problem can be presented by materials exhibiting excessively coarse primary fragment distributions, y values > 0.9. Breakage of this type of material produces size distributions that give poor packing efficiencies even at long grind times, thus hindering the normal rheological transformation presented earlier. A final condition that can cause the occurrance of region B to be small or zero is when the media void volume filling of slurry is < 100%. Figures 5a and 5b use previously published data of this study to demonstrate the influence of solids loading and weight percent solids on the net production of taconite ore in a laboratory batch ball mill. In particular, Fig. 5a illustrates several trends not generally recognized until this study. They include these two: • Increased slurry density allows for increased solids loading before passing through the maximum in the net production curve where the fall-off is due to non first order breakage. • The use of rheology control chemicals such as GA-4272 allows this trend to be extended to higher slurry loadings with an increase in net production over any previous condition by keeping grinding first order. Figure 5b shows the same data as Fig. 5a plotted for constant weight loadings. These various figures of net production versus percent solids show the location and extent of the regions A, B, and C presented in Part 1 as a function of solids loading. It is obvious that the rheological transformation pattern described earlier does not hold true for slurry loadings corresponding to less than the void volume of the media, which is also a region of

Jan 1, 1983

By Malcolm J. A. Swallow

In June 1997, Mount Polley Mining Corp., a division of Imperial Metals Corp., completed an 18-kt/d (20,000 stpd) gold-copper mine in central British Columbia. Fast-track construction and extensive off-site construction were combined with an aggressive project schedule to build the mine in one year. This was three months ahead of schedule and 7% below the original budgeted cost of C$123.5 million (US$86 million). The Mount Polley Mine was the first of a crop of new open-pit mines in British Columbia. The Mount Polley deposit has been explored since the mid-1960s when it was known as the Cariboo Bell property. From 1988, Imperial Metals completed drilling and metallurgical testing on the deposit. This led to a positive feasibility study and the issuing of a Mine Development Certificate in 1991. The project was later shelved due to falling metals prices and the strength of the Canadian dollar. In late 1994, Imperial merged with Bethlehem Resources Corp., bringing a strong project team to the company. The deposit was then re-evaluated, ore reserves revised and additional metallurgical testing was performed. The result was a revised project plan. It indi¬cated that a financially feasible gold-copper mining op¬eration could be constructed at Mount Polley operat¬ing at 18 kt/d (20,000 stpd). In May 1996, an agreement was reached with Sumitomo to jointly fund and develop the project. By early 1996, final plans and permits were being prepared for construction, scheduled to begin in June 1.996. On-site work commenced on May 28th, 1996 following lifting of the road restrictions. Mount Polley Mining acted as its own project manager. Construction was completed in June 1997. Plant runup took place in July and full production began in August. Through the use of off-site construction and the careful selection of contractors who cooperated to complete some 12 major contracts, it was possible to reduce the mine and mill construction time from 16 to 13 months. Off-site construction for all possible aspects of the work and the use of used milling equipment allowed significant savings in overall capital costs of construction. Total costs for the mine, mill and tailings dam equated to C$6,100/t/d of throughput. The mill design incorporates several innovative features that will aid maintenance and allow flexibility of operation. The operation is designed as a closed system and meets the latest environmental standards. Concentrate is shipped by truck to the port of Vancouver. Revenues for the mine come from the mine&apos;s production of about 3.1 t/a (100,000 oz/year) of gold and 10.8 kt/a (24 million lbs/year) of copper during the first four years of the 12-year mine life.

Jan 1, 1998

By R. R. Oder, R. J. Gray

We report measurements of the effects of rank, type, and grade of coal on grinding energy and grinding media consumption in a Pitt Mill. The rank of coal has a greater effect on grinding energy than does coal type. The effect of coal grade on grinding energy depends upon the relative grindability of the mineral matter and of the hydrocarbon components of the coal. The mineral matter is primarily responsible for abrasive media wear in a Pitt Mill. The widely recognized HGI test for coal breakage may yield misleading results, particularly for multiple-banded bright and dull coal where the screening procedure used in sample preparation favors the testing of the hardest coal fraction. The work reported here indicates that the Pitt Mill, which grinds the whole coal in a single stage and which yields easily interpreted results, is an excellent candidate for testing coal pulverization characteristics including energy requirements and equipment wear. The measurements of the effects of coal characteristics on grinding energy and on media consumption have been used to estimate the cost of fine grinding in a Pitt Mill. The results, based on a conceptual level evaluation of grinding at the rate of 1 TPH, show that there are trade-offs possible between the cost of energy and cost of media consumption which involve choice of grinding media size. The use of fine-size grinding media to lower grinding energy and to increase throughput can be deceptive because the total cost of grinding must consider both energy and media consumption.

Jan 1, 1992