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By A. L. Mular

"IntroductionThe mining industry has made slow but steady progress toward the automation of mineral processing plants over the last 35 years. The stimulus to automate is associated with important economic and hidden benefits, so that today, the latest in instrumentation and technology is being employed.This chapter assesses the status of instrumentation justified economically, defines levels of control and certain fundamentals of feedback control loops, describes typical sensors employed in concentrator circuits, summarizes control strategies employed therein, and briefly highlights trends in integrated plant control.Current Status of AutomationAs early as 1948, the mining industry began to show interest in automation, probably because of its close association with the petroleum industry. By 1958, some regulatory control had been achieved in various sections of a concentrator, particularly in crushing circuits. Successes were attributed primarily to the reliability of sensors employed in the corresponding control scheme. By 1961, computers had appeared on the scene, initially, for data logging purposes.From 1958 to 1965, very few new control strategies had been implemented. Guidelines in vogue at this time appeared to be:- Install proven automation, unless a good instrumentation department exists at the concentrator.- Do not bite off too much instrumentation at one time.- Do not purchase on price alone; instrumentation must function in the concentrator environment.- Do not purchase automatic control equipment to be fashionable.Such guidelines were employed1 because of a lack of: (a) quantitative process knowledge, (b) reliable primary sensors, (c) initiative to try something new, (d) trained personnel, and (e) understanding that older concentrators could justify automation in many situations."

Jan 1, 1989

By B. J. Arnold, F. F. Aplan

"The captive bubble and sessile drop contact angle techniques have been used to evaluate the hydrophobicity of petrographically identified coal macerals. The magnitude of the contact angle on vitrinite is shown to be a function of coal rank ( %C, %0 and % vitrinite reflectance), increasing with increasing rank to a maximum for low volatile bituminous rank coals then decreasing for anthracite. The least data scatter is found when vitrinite reflectance is used as a rank parameter. Contact angles measured or estimated for other discrete macerals, indicate a variation in hydrophobicity based on coal type. These data correlate well with known flotation results. Coal is a heterogeneous mixture of plant remains and associated minerals that have undergone chemical and physical changes by biological and geological processes. These plant remains form coal macerals, discrete organic entities in the coal which have characteristic chemical and physical properties. As the differing chemical nature of each maceral will affect its behaviour in processes such as carbonization the separation of macerals during coal preparation may provide an important asset in coal utilization.Separation of some of the macerals based on specific gravity has been achieved in the laboratory, and by Sun and Cohen , in a study of two Pennsylvania bituminous coal preparation plants. It was indicated that the lighter exinites report almost exclusively to the clean coal while the denser inertinites are concentrated in the higher gravity fractions. The Longwy-Burstlein process6 took advantage of the difference in hardness or friability of the macerals by using impact breakage followed by screening to remove the more friable vitrain,clarain and fusain. The remaining durain, often high in spore content, was then recrushed. The re-mixed fractions produced the optimum size for obtaining the best coke strength for a particular coal.Separation or concentration of rnacerals by froth flotation largely depends on differences in the hydrophobicity of the macerals. Sun and Cohen 5 . indicated an order of decreasing floatability as exinite, vitrinite and inertinite, while a Russian study (as cited by Brown 7 ) showed that vitrains floated more readily than the duller types. More recently, Hirt and Aplan8 showed that 'pseudovitrinite' floats somewhat faster than does vitrinite while fusinite floats very slowly. Olson and A plan found that vitrinite floated easily while substantial quantities of the non-floated coal contained fusinitic material. Burdon noted that vitrinite is more floatable than semifusinite, and Sarker et a/. 11 have shown that vitrinite is selectively concentrated by froth flotation and oil agglomeration. Sun indicated that flotation, based on his calculated floatability index, should decrease in the order clarain, vitrain, fusain and durain; while Horsley and Smith 13 indicated the decrease to be: vitrain, clarain, durain and fusain. These differences can be attributed to the specific coals used in each study and to the individual macerals composing each coal type within a given coal."

Jan 1, 1989

Provide mine personnel with a convenient means of determining the optimum installation geometry of roof trusses in bedded roof. Background Mine operators frequently ask the following questions concerning roof truss installations: 1. Where should the bearing plates or brackets be placed or where should the inclined holes be started? 2. At what angle should the holes for the inclined chords or angle holts be drilled? 3. What tension should be put into the inclined chords so as to achieve the most effective results for the effort and expense? Truss hardware manufacturers have responded to these questions with answers based on what seems to work best. While designs based on past experience are frequently reliable, there was a need for a rational approach that could accommodate a variety of conditions. The Bureau has developed a procedure that enables mine operators to optimize roof truss installations in bedded roof.

Jan 1, 1989

By Betty L. Gibbs

Mine planning and scheduling done with computer programs is possible only after extensive work has been completed to define the geologic extent of an ore body and an overall mine design. Mine planning methodology on computer can vary from computerized engineering tasks through systems which completely optimize a plan with little input from an engineer during the actual scheduling process. However, a strong emphasis is placed on producing practical mining sequences which may be "best" rather than "optimum". This section offers several viewpoints on the planning process using computers. Interactive planning and scheduling systems are attractive for many companies because they are easy for engineers to understand, and the results are acceptable. The engineer develops an open pit mine sequence by trial and error using interactive graphics. The results of this approach can usually be improved. A computer program can help an engineer make many trials in a short time period and provide a higher level of confidence that a "best" plan has been generated.

Jan 1, 1989

By S. McTavish

Since mechanically suitable grinding mills were available in 1908, mill operators have been searching for the best material and design with which to line the mills. These liner systems had to fulfill several criteria. First and foremost they had to last long enough in an operating mill to protect the mill structure from the abrasive wear of the ore and the grinding media. In addition they had to be able to efficiently transfer the energy from the mill to the mill charge in order to facilitate the grinding process. The liners also had to be easily attached to the mill when new and removed from the mill when worn out. The first liners used were generally made of cast white iron, Hadfield manganese, unalloyed high carbon steel and composites of steel and concrete or cemented in flint blocks or natural stone. 1

Jan 1, 1989

By C. J. Veal, C. I. House, S. R. Bellamy

INTRODUCTION Coal gold agglomeration (CGA) is a new concept in gold processing. It involves the recovery of oleophilic (oil-loving) gold grains by causing them to adhere selectively to the surfaces of coal oil agglomerates. Under optimum conditions very high recoveries of gold. can be achieved. The agglomerates are recycled to build up gold loading to typically 1000 - 5000 g/tonne. Bullion can be produced by burning the loaded agglomerates with subsequent recovery of gold from the highly loaded ash either by dissolution/electrowinning or smelting. A wide-ranging programme of laboratory testwork has been performed to investigate the response of different ore types to the process. This has shown CGA to be most applicable to ores in which the gold is liberated or easily liberatable, relatively coarse (ca 500 µm) and fine (<5 µm) gold grains being recoverable with equal ease. The best opportunities for CGA seem to be: • placers containing fine gold not recoverable by gravity concentration • low grade gravity concentrates to avoid multiple magnetic/gravity separation stages and amalgamation • free milling ores where the gold can be liberated at a reasonable ore particle size eg >d80 50 µm • tailings past or present from gravity concentration plants. CGA testwork ranging in scale from laboratory to pilot plant on tailings produced by gravity concentration in the 19th century Australian gold rush has already been reported (Bonney, C F, 1988; House, C I, Townsend, I C and Veal, C J, 1988). An economic appraisal for a Canadian free milling gold ore has also been published (Buckley, S A, House, C I and Townsend, I G, 1989). This paper describes results of a detailed laboratory programme to assess CGA as a means of recovering gold from a beach sand deposit in New Zealand. The area from which the samples were taken consists of quarternary interglacial marine, lagoonal, fluvial and eolian deposits. The heavy minerals present are predominantly ilmenite and garnet and small concentrations of gold. In certain locations the sands and gravels are consolidated by a limonitic cement. In the 19th Century some parts of the deposit were mined and the ore processed using stamp milling and gravity concentration. Gold recoveries were low, probably below 50%, due primarily to the fine size of the gold grains, the flaky nature of the gold and also imperfect liberation by the stamp mills. Our laboratory testwork was divided into three parts: i) Preliminary studies on an initial sample to assess the feasibility of treating the ore by CGA ii) Extended studies on further samples from various locations in the deposit to confirm predictions from (i) iii) Conventional gold recovery techniques. EXPERIMENTAL PROCEDURES All CGA tests were performed by agitating a slurry of ore in water in a 0.6 dm3 pyrex beaker fitted with four vertical baffles. When required, potassium amyl xanthate collector was added and the slurry conditioned for 5 min. Coal (ex BP Coal Australia&apos;s Tahmoor Colliery) previously ground to a d80 50 µm was added followed by a commercial grade of gas oil (density 840 kg/m3). The agglomeration time was 30 min. The agglomerates were separated from the tailings by screening, both were then dried and analysed for gold by fire assay. This test procedure provides a reasonably accurate estimate of gold recoveries that might be expected from our 1 tph pilot plant. Ore samples were ground either dry in an agate ring disc mill or wet in a steel batch tumbling

Jan 1, 1989

By E. Alaphia Wright

An accepted approach to open pit mining sequence planning is one of computer aided planning where: the planner works out his mining sequences by trial and error using interactive graphics; This approach gives results that could in- variably be improved given more time for the planning exercise. Increased emphasis on interactive planning reveals a need for simple formalized procedures for improving mining sequence planning practices. This paper discusses a case study in which Dynamic Programming (DP) is used in mining sequence planning. It is shown that DP can be applied to identify various intermediate, pit boundaries. These are then arranged in descending order of economic values and their material contents analysed with respect to tonnages, grades, production requirements, and equipment capacities. Those increments which best satisfy defined conditions are selected and the others discarded. The mining sequence is thus built up as a chain of best- economic valued mining increments satisfying given production requirements and constraints.

Jan 1, 1989

By R. L. King

An expert system is a computer system that performs tasks that are considered by most people to require human expertise. Since their functioning relies heavily on large bodies of knowledge, they are often referred to as knowledge-based systems. These systems can contain the skill and knowledge of an individual human expert or the combined knowledge of a group of experts. The knowledge embodied in an expert system can consist of specific facts, HSHA regulations, heuristics or rules-of-thumb used by the expert, and databases. The purpose of this paper will be to introduce the participants in the ?Increase Competitiveness Thru Innovation? session to expert systems. This introduction will enable the participants to better appreciate the balance of the papers to be presented in this session. These papers will deal with specific mining problems and a proposed innovative solution. This paper will discuss such topics as problem-solving paradigms, rules of inference, knowledge representation, and the structure of expert systems.

Jan 1, 1989

By John E. Pahlman

The U.S. Bureau of Mines has continued its investigation of chemical additive drilling fluids that produce a zero surface charge (ZSC) on a rock to better define the boundary conditions of the phenomenon. Laboratory drilling tests were performed on Minnesota taconite, Tennessee marble, and Sioux Quartzite using diamond-impregnated coring bits or tungsten carbide spade bits. Drilling fluids were made from chemical additives such as inorganic salts, cationic surfactants, cationic polymers, acids, or nonionic polymers in either distilled, deionized water (DDIW), mine pond water, mine well water, or tap water. These additives were tested below, at, and above their respective ZSC concentrations. Penetration and bit life improvements obtained with ZSC concentration solutions ranged from 88 and 56 pct, respectively, to over 650 pct and over 400 pct, respectively, for the additives tested. Polyethylene oxide (PEO) was found to be the best additive for ZSC-controlled drilling because a continuous range of PEO solution concentrations produced the enhanced ZSC-controlled drilling performance.

Jan 1, 1989

By P. D. Laverty

The U.S. Bureau of Mines treated mixed and contaminated super alloy scrap by pyrometallurgical and hydrometallurgical methods to separate and recover metal values. Best results were obtained by leaching Zn-treated or atomized scrap with HCI-02 at 9SO C and 50 psig 02&apos; This resulted in dissolving approxi-mately 98 pet of the AI, Co, Cr, Cu, Fe, Mn, Ni, and Zn while rejecting over 98 pct of the Mo, Nb, Ta, Ti, W, and Zr as an insoluble refractory residue. Chlorine was successfully substituted for HCI to leach Zn-treated scrap but was unsuccessful for leaching atomized scrap. The leaching solution was treated by pH adjustment and hydrothermal precipitation at 2000 C for 4 h to remove AI, Cr, Fe, and other contaminants as a filterable precipitate. Recovery of Co and Ni would be accomplished by solvent extraction and e1ectrowinning. Chromium recovery as a ferroalloy was demonstrated.

Jan 1, 1989

By A. M. Blount

In 1983, a final ruling of the U.S. Occupational Health and Safety Administration (OSHA) specified that chemicals containing >0.1 % of a carcinogenic substance or > 1 % of a substance hazardous to health must be labelled. Amphiboles of the fibrous or asbestos-type must be quantified at levels of 0. I % whereas the status of non-fibrous types is not finally determined at this time. Tremolite and actinolite can possibly be measured to 0.1 % by XRD, but anthophyllite cannot be so measured. Recently the International Agency for Cancer Research (IARC) has designated quartz as a suspected carcinogen, and its quantity must be measured to the 0.1 % level. Thus, it has become Increasingly important to be able to detect and quantify minerals in bulk samples down to low or "trace" levels. It has been found that minerals needing to be quantified can often be concentrated by microcentrifuge and counted by means of polarized-light microscopy. For example, amphiboles in talc samples can be separated by using a heavy liquid (thallium formate-malonate) of density 2.810. Talc floats, and amphiboles sink. After centrifuging, the denser particles are removed from the bottom with a Pasteur pipette and placed in a clean centrifuge tube. Distilled water is added to the tube containing the denser minerals and the tube centrifuged again. After several washings with distilled water, the sample is removed along with a drop of water to a clean glass slide. A drop of 1 .584 refractive index liquid and a cover glass are placed over the dried sample. The slide is scanned for particles greater than 5 um in length and having an aspect ratio greater than 1:3. These are checked for refractive index, sign of elongation and extinction angle to differentiate them from talc particles on edge and other minerals. Amphiboles have all indices greater than 1.584, and talc has all indices equal to or lower than this value. Alternately the heavies at the bottom of the centrifuge tube can be filtered through a nucleopore filter. After the filter has been washed with distilled water and dried, a 1.584 refractive index liquid renders the filter transparent in two of the four extinction positions so that a count of amphibole particles can be made directly on the filter. The refractive index-density method which is so useful for talc powders should be equally useful for other nonsoluble industrial minerals.

Jan 1, 1989

It is a well known fact that the best method of rock reinforcement is to avoid physical ground support. This goal can be achieved by means of: - Increasing the rock structural strength - Geological definition and forecasting. - Sequencing the excavation - Smooth blasting In many cases however, extra ground support is needed in order to maintain safe working conditions. In an open pit, the structural strength of rock cannot be increased by lowering the pit walls due to economical reasons. Economy demands steeper walls and less waste to be hauled to the dump. The geological conditions can also often be extremely poor and unpredictable. For excavation sequencing and smooth blasting there are good possi- bilities, such as cleaning of the pit walls after the main excavation, arul- ing of long small diameter straight inclined holes, use of smooth blasting techniques and lightened explosives.

Jan 1, 1989

"This generic study has been developed in response to the frequently asked question, ""That is the best treatment route for refractory gold ore?"". The study has developed operating and capital costs for many potential treatment routes and they are presented here in a generic form. All costs are in relation to the value of 1.00 given to the base case, a 1000 mtpd bio-oxidation plant, operating at 25% solids and 72 hours residence time.There is unfortunately no simple answer to the above quest ions. The treatment of refractory ores is a complex subject, with many variables, such as mineralogy, chemistry, environment and tonnage. However this study does offer information and relative numbers that we believe will be of assistance to gold metallurgists in obtaining a ""first shot"" at estimating costs.Equally, this technology is a rapidly evolving one. As more plants are commissioned, and test programs completed, better data is constantly being generated. For a generic study of this type to remain useful it will need regular updating and adjustment."

Jan 1, 1989

By W. K. Connor, O&apos

The U.S. Bureau of Mines investigated the Sheep Creek prospect in the north-central Alaska Range to evaluate the Pb, Zn, Ag, and Sn resources. The studies included field investigations, mineralogical characterization, and concentration tests. Mineralization at Sheep Creek consists of disseminated, banded, and massive pyrite, sphalerite, and galena. Tin occurs as grains of cassiterite (Sn02) in the quartz-mica matrix and within pyrite and sphalerite grains, and Ag occurs in solid solution within the galena. Three bulk surface samples and two sets of core samples were collected at the Sheep Creek prospect for this study. Gravity concentration yielded products with up to 30 pct Sn, and the best result was a product that recovered 52 pct of the Sn at a grade of 22 pct Sn. Bulk sulfide flotation yielded products that recovered up to 89 pct of the Zn, 73 pct of the Pb, and 75 pct of the Ag, but grades were relatively low; selective flotation would significantly increase the grades.

Jan 1, 1989

By Roland Loewen

A review of the aqua regia method for the recovery of gold from jewelry shop wastes is given. Also outlined is a specialized procedure for the recovery of the silver that occurs in these wastes. A classic method for the recovery of platinum and palladium is, given. There is a brief review of other gold and silver refining methods. Environmental conditions are discussed and a few perhaps, new and useful processes ,are mentioned. References are by author or editor and are listed alphabetically. Suggestions for further information in the field of precious metals, are given in a final list.

Jan 1, 1989

By M. J. Meulendyke

Within the mineral processing Industry. a range of grinding conditions exists which include semi-autogenous grinding (SAG), rod milling. and conventional ball milling. Each of these mill environments presents a unique environment for grinding media, requiring the application of specific physical and chemical properties for optimum grinding media performance. The environments are characterized by varying degrees and combinations of abrasive, corrosive, and impact wear. An extensive test program has been conducted to determine the extent wear rates vary between these different applications. Test results are related to production results, ball size, and mill operating conditions.

Jan 1, 1989

By V. Srajer

This paper deals with the problem of efficiently selecting the most appropriate loading and hauling equipment for surface mines. Problems and factors that affect productivity and operating cost are identified. During the course of collecting the information for this study, mining companies and equipment manufacturers were interviewed to define their selection practices. The majority of mining companies prepare general specifications and ask the suppliers to mate those specifications to their product line. The mining company then evaluates various proposals on the basis of the best available deal. The following &apos;mines from western Canada provided information for this study: Coal Mountain, Line Creek, Quintette, Westar&apos;s Harmer mine from the Province of British Columbia, and Coal Valley, Gregg River, Highvale, Luscar from the Province of Alberta.

Jan 1, 1989

By Todd B. Harrington, Dennis M. Lydon, Walter B. Sudweeks

Heavy ANFO products made by mixing ANFO (or AN prills) with a relatively small amount of emulsion (typically 15-30 percent) have become very popular in recent years. They are less expensive than straight, sensitized emulsion and maintain the good shooting characteristics of ANFO but have higher bulk strengths. Their use, however, is generally limited to larger diameter, dry boreholes. Smaller diameter, wet-hole quarries have not been able to widely utilize this low cost bulk product.

Jan 1, 1989

By D. C. Durgin

Located near Ely, Nevada, the Robinson Mining District has produced gold since 1868 from the Ruth? porphyry copper system. Gold occurs in at least seven different deposit types, including porphyry copper ore, silica-pyrite replacements along favorable horizons, fault controlled silica bodies, intrusive breccias, stockworks in bornfels and in retrograde skarns. In all of these the primary controls localizing gold ore are structural .preparation and permeable hosts. Indicator elements associated with gold are Ag, As, Te, and to a lesser extent Sb and Tl. Pb, Zn and F are peripheral to gold or with it depending upon position within the system. The district has produced 87,000 kg (2.8 million ounces) of gold thus far. An additional 19,300 kg (620,000 ounces) of reserves are established, and well over 31,100 kg (1,000,000 ounces) of reserves are anticipated. I

Jan 1, 1989

By J. Avraamides

Some operational difficulties have been encountered in the transfer of carbon-in-pulp technology to Australian conditions. Many gold mines? are located in remote and arid areas where the available ground water is highly saline. There is a gradual depletion of free milling oxidised gold ore reserves resulting in the necessity to treat deeper primary sulphide .deposits which can be refractory. Some near-surface ore-bodies and tailing dumps are contaminated by natural vegetative matter which can interfere with the adsorption of gold from the cyanided pulp,. Advances which have been? made include potable water conservation in elution, a new type of regeneration kiln, inter-stage screens, alternative leaching systems transportable carbon-in-pulp plants and efficient open-cut mining methods. Academic and applied research is well supported by government and industry and the results are normally rapidly disseminated and put into practice by the gold mining industry.

Jan 1, 1989