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By Fathi Habashi

ION FLOTATION was introduced by F. Sebba in 1959. This technique permits the recovery of metal ions from solution through the introduction of an appropriate cationic or anionic surface-active agent into the solution. The surface-active agent, called a collector, is added in stoichiometric amounts in relation to the metal content of the solution. It reacts with the metal ion to form an insoluble soap, and is levitated to the surface by gentle bubbling of air to form a foam which collapses into an insoluble scum. The optimum conditions for the ion flotation with soaps of C10-16, fatty acids are shown in Table 15-1. Baarson and Ray (1965) found that once the metal was completely precipitated as the hydroxide, as little as one hundredth the stoichiometric amount of the collector was required to cause complete flotation of the [ ]

Jan 1, 1970

By G. Chi, M. C. Fuerstenau, R. C. Bradt, A. Ghosh

Technical Papers, Mining Engineering Vol. 49, No. 12, pp. 70-75, Discussion by McIvor, R. and Fink, R. P. We must challenge what we consider to be dubious statements in the above-mentioned article. Much of what is reported was published elsewhere over a year ago (i.e., in their reference: Chi, G., Fuerstenau, M.C., Brandt, R.C. and Ghosh, A., 1996, "Improved comminution efficiency through controlled blasting during mining," International Journal of Min¬eral Processing, Vol. 47, p. 93). There is nothing unusual or significant about a 10% or more difference in grindability between two materials of apparently similar origin and other attributes (such as chemical assays, hardness and specific gravity). Consider also that: • the materials came from different locations in the pit; • only two sets of four blastholes were considered per blast: • spatial orientation with respect to the face and other blastholes was also altered; • only two pairs of data were collected; • the impact of sampling error and bias, • both of which could be significant in this case; and • normal experimental error in parameter measurement and calculation. It is clearly inappropriate to draw any conclusions with regard to cause-and-effect relationships without first presenting a thorough error analysis. A substantial number of randomly selected test and control areas might provide meaningful statistical evidence of the claimed effect of the variables being studied. Even that appears doubtful, however, because Figures 3 and 4 show about as much variation in crushing energy for the two "regular" conditions vs. regular and altered conditions. On a separate, but also important issue, broad experience with autogenous/semiautogenous milling has shown that smaller or weaker feed material results in reduced tonnage. ? Reply Thank you for the opportunity to respond to the comments of Messrs. Mclvor ant Fink. The points raised are addressed in the order of presentation. As the commenters noted, there are parallels between this article and a previous manuscript. The reason for these complementary articles is that the audiences for the two publications are different, although there is obviously some overlap. We believe that the different audiences each profit from the technical aspects of the two manuscripts. We are not in agreement that a grindability difference of 10% is insignificant. The laboratory work is corroborated by data obtained at plant scale, which confirms that an increase in throughput of the same magnitude (between 5% and 10%) is achieved when the blasting pattern is modified, i.e., reduced from 9 to 8 m (29.5 to 26.2 ft). This increase in throughput translates into substantial power savings for ore grinding. No differences in appearance in the rock were noted between the two adjacent areas. Rock in both areas belonged to the same rock mass/block, which was quite homogeneous and without geological features in between that could create drastic differences in comminution characteristics. Given the resources, we would have sampled from all of the holes. Additional data would certainly have been desirable, as would have been additional funding to expand the study. Their comment, "spatial orientation with respect to the face and other blastholes was also altered," is not clear. Figure 1 does not show the initiation sequence. Without the initiation sequence, we do not understand how the commenters concluded that spatial orientation with respect to the face and other blastholes was altered. The commenters may be trying to make the point that the four-hole configurations with respect to the original bench face are not exactly the same. In fragmenting the

Jan 1, 1999

It doesn't take any special acumen to note that US engineering and construction firms are busier than ever with mining projects in this mineral conscious world. This special issue of Mining Engineering takes a look at E&C companies-their expanding role, their special capabilities, and their challenges and problems. ME sent a questionnaire to a representative sample of E&C firms. Information was sought about officers, ownership, and amount of business. ME also asked about types of projects and commodities dealt in as well as market distribution and forms of payment. The accomnanying charts show how the 17 responding E&C firms answered these and other questions. Other ME survey questions, and a review of company literature, yielded information about financing techniques and special challenges for E&C firms. Insights were gained concerning government impacts, customer-generated problems, and promising areas for development-including synfuels-in the minerals industry. These areas are discussed later in this article. Accompanying articles in the special issue discuss the E&C company's role in mining projects, research and development in E&C firms, one approach to contract work used by an E&C firm, and an E&C project case study.

Jan 10, 1980

By Rodolfo V. de la Cruz

The in-situ deformability of a highly jointed basalt rock mass was determined by two distinctly different methods: one, by the NX- borehole jack method where the displacements of opposing curved platens were related to the applied hydraulic pressures, and; two, by the modified Goodman jack method where the tangential strains on the borehole walls were related to the induced tangential stresses. The modulus obtained by the modified Goodman jack method were much higher than those obtained by the NX- borehole jack method. To explain the discrepancy, the influence of fractures and test variables such as depth, orientation, hole number and applied pressure on the calculated modulus of the rock mass were analyzed by factorial analysis and it was found that the orientations and depths of measurement has statistically significant effects. The in-situ deformability values obtained by non-linear regression analysis were also found comparable with other measurements and empirically predicted values for the basalt rock mass.

Jan 1, 1982

By M. Paul

"INTRODUCTION In 1990, the uranium production of SDAG Wismut in East Germany was stopped. 45 years of inconsiderate destructive exploitation had scarred both landscape and nature. With the “Wismut law” which came into force in December 1991, the Federal Republic took responsibility of the remediation of the legacies from former uranium ore mining and milling in Saxony and Thuringia. A unique opportunity turned out in order to remove environmental damage, to enhance the life quality of the local population, to shape new landscapes and to advance the economic restart. (1) The results of nearly 25 years of remediation activity are evidence that this opportunity was used at all former SDAG (Sowjetisch-Deutsche Aktien-gesellschaft) Wismut sites. The reference project “Wismut“ has evolved into a joint project of all parties involved and radiates beyond the affected regions. This is particularly reflected by the Know-how transfer starting from Wismut and the increasing activities for third parties abroad.WATER MANAGEMENT AT WISMUT REHABILITATION Within the Wismut project, water management and conventional technical water treatment are by far the most cost-intensive long-term tasks of the remediation programme. Over the medium term, there is no viable alternative to the operation of active systems to catch and treat contaminated mine waters at the Ronneburg, Schlema, Königstein, Pöhla, Seelingstädt and Helmsdorf sites. Based on the status quo, this paper outlines the key issues of the Wismut GmbH water management strategy over the medium and long term. It is focused primarily on achieving protection goals for potentially impacted water bodies in the surroundings of Wismut sites and on optimising associated remediation expenditures as well as on creating the prerequisites for achieving low post-remedial care and maintenance or walk-away system status over the long term. (2) The applied mining practices during the active SDAG Wismut period have not been up to the high standards of modern mining operations and their requirement to leave former production premises and deposited residues in a condition that was not only physically but also chemically stable. Rather the emission of mine and seepage waters laden with high acidity, metal and salt content is a frequent corollary and difficult legacy of historic mining and processing operations (4, 5). As a consequence, the control or mitigation, respectively, of environmental impacts associated with contaminated mine waters ties up enormous public resources worldwide (6, 7)."

Jan 1, 2016

By Bruce D. McCellan

INTRODUCTION The gathering arm loader was patented by Joseph Joy in 1916. This unique machine motion became the standard of the industry and is used on underground loaders and continuous miners throughout the world. Figs. 1 and 2 show two types of gathering arm loaders. The gathering arm loader is a crawler-mounted ma¬chine primarily used for loading coal in a conventional mining system. Other uses also have been found for this machine in noncoal mines and other industries. MECHANICAL ASSEMBLIES The front or loading end of the machine, commonly called the head, consists of a tapered structural frame, hinged at the rear to the chassis and riding the ground at the front. Two bevel-gear cases are mounted in the head. They are connected by a timing shaft, which also acts as a conveyor drive shaft on most models. An eccentric disk attached to each gear case drives the gathering arms. The rear of the arms are attached to a swivel assembly. The motion derived from this arrange¬ment gathers loose material from the ground directly in front of the machine and loads it into the conveyor. Power for this drive is provided by two electric motors mounted under the gathering head. A speed reducer attached to each motor provides a gathering arm speed of approximately 5.2 rad/s (50 rpm). A torque limiting clutch is provided on each reducer to guard the system against sudden impact loading and continual overloading of the motors. The conveyor consists of an open box section through which a universal-type chain carries the ma¬terial to the rear of the machine, where it is discharged into the haulage vehicle. The chain is driven by a sprocket located on the timing shaft between the two gathering-head bevel-gear cases. On some large loader designs, primarily for noncoal applications, the chain is driven by a separate gear motor at the discharge end of the conveyor. However, the system driven by the timing shaft is used most widely. The conveyor frame is hinged just to the rear of the gathering head. It can be raised or lowered by hydraulic cylinders to the desired discharge height. The conveyor frame also is designed to pivot 0.79 rad (45°) either side of the center, just to the rear of the chassis. This feature increases the maneuverability of the ma¬chine and makes it possible to discharge material into the haulage system when in tight quarters such as when turning a crosscut. This function also is controlled by a hydraulic cylinder. The chassis of the loader consists of two crawler frames and cross members, holding the frames in position. The frames are either welded or cast as¬semblies. Replaceable wear runners are provided on the frames to take the wear generated by the crawler chain. An adjustable take-up idler keeps the crawler chain in proper adjustment for maximum life. The drive system for the crawler chains varies with the manufacturer, but it generally is one of two basic designs. One system provides a traction motor for each side of the machine, coupled to a speed reducer. A spur and bevel-gear train transfers power to the sprocket, which drives the crawler chain. Directional control is achieved by disk-type clutches. The other system also uses a motor for each crawler, coupled to a speed reducer. A roller chain drives the sprocket, which in

Jan 1, 1982

By Changhong Huang, Y. J. Wang

This paper describes the formulation of mine ventilation network optimization as a standard nonlinear programming problem and the solution of the problem using the generalized reduced gradient (GRG) method. In general, the GRG method is regarded as one of the most effective and efficient methods for solving large-scale nonlinear programming problems. The variables considered in the network optimization include air quantity, regulator pressure loss, and fan pressure for each branch. Each of the variables is constrained by lower and upper bounds. This facilitates proper assignments of ventilation requirements and provides a great flexibility to include various situations encountered in mine ventilation. The objective of the network optimization is to minimize the overall air power consumed by or supplied to the network; whereas the constraints are, in addition to the variable bounds, the two fundamental network laws: Kirchhoffs voltage law and Kirchhoffs current law. A computer program has been developed based on the GRG method for the network optimization. An example of network with 18 branches, 8 regulators and 3 fans is included for illustration.

Jan 1, 1993

By M. L. Young, M. S. Klima

Hindered-settling columns are used throughout the minerals industry for both size and density separations. In this study, the performance of a continuous laboratory-scale hindered-settling column was evaluated for various particle systems. The test variables included solids composition, water rate and column height. Column performance was evaluated using partition curves and corresponding performance parameters.

Jan 1, 2000

By Donald E. Ralston, V. Rajaram, Michael N. Greeley, John W. Peters, Terry J. Laverty, F. S. Kendorski, Peter G. Chamberlain, J. Brent Hiskey, William C. Larson

Preliminary figures from the US Mine Safety and Health Administration show 801 underground metal mines and 490 surface metal mines, at the end of 1981, up from 663 and 383, respectively, in 1980. Preliminary MASH figures also show 207 underground nonmetal mines and 1070 surface nonmetal mines at the end of 1981, compared with 167 and 1092, respectively, last year. In addition, preliminary MASH figures show there were 14,676 mineral extraction operations on the surface and 3845 underground at the end of 1981, versus 14,951 and 3675, respectively, for last year. The number of mines, then, in three of four categories increased significantly, while having a lower net total of just 2% in the other category. And the number of underground extraction operations increased nearly 5% for the year, while surface operations were down less than 2%. But despite operations totals that mostly held their own or showed significant increases, it was a tough year for the nation's mining industry. The nonferrous segment of the industry had a bad year and 1982 may be no better, according to Business Week (Jan. 11, 1982), which adds that a turnaround is impossible unless the general economy recovers. Even then, with diminishing markets and overcapacity plaguing its products, the mining industry's condition "will be far from robust," the magazine says. The basic problem is that nonferrous metals-aluminum, copper, lead, molybdenum, nickel, and zinc-are tied to the most troubled sectors of the US economy. As recently as 1979, for example, more than 40% of all aluminum shipments went to the construction and transportation industries. Last year, these markets were down to 35% of aluminum shipments. And molybdenum sales to steel producers dropped by nearly 15% last year. Moreover, nonferrous products ended 1981 at disastrously low prices. Molybdenum oxide sold at $8.77/kg ($3.98/lb), half its $17.63/kg ($8/lb) price earlier in 1981 and aluminum was listed at $1.08/kg ($0.49/lb), compared with $1.54/kg ($0.70/lb) a year earlier. The Comex price of copper was $1.54/ kg ($0.70/lb), down from $3.11/kg ($1.41/lb) in early 1980, in real terms the worst copper price in 30 years. Lower prices led operators to cut their output and increase layoffs, shut down mines, and defer capital spending projects. In December, Duval Corp., a Pennzoil Co. subsidiary, shut all four of its US copper mines until March. Aluminum Co. of America had laid off 7.3% of its work force, shut nine of its 38 smelting lines, and entered 1982 operating at only 68% of capacity. Inco Ltd., the Free World's largest nickel producer, lost money in 1981, an estimated $11 million, for the first time since 1932. And, like Falconbridge Nickel Mines and other major nickel producers, had cut back on nickel production and was operating at only 63% of capacity (Forbes, Feb. 15, 1982). Amax, the country's largest molybdenum producer, had cut production by 20%. And Hanna Mining Co. sold lead for $5.07/kg ($2.30/lb), about $1.54/kg ($0.70/lb) below what it cost to produce. Another problem is that copper companies in the next few years must come to terms with problems of aging smelters. Some companies are simply closing them. In 1980, Anaconda closed its 75-year-old Butte, MT, smelter and is shipping concentrates to Japan, rather than modifying the smelter to meet environmental requirements. And oth Phelps Dodge Corp. and Asarco Inc. have antiquated smelters on which they will have to make decisions. The most troubled metal in 1982, Business Week said, will be molybdenum, with no real recovery in sight until the mid-1980s. High molybdenum prices in the late 1970s, caused in part by shortages, triggered overexpansion that will take years for the US market to soak up. On a more upbeat note, 1981 mergers might help stabilize the industry by providing funds for mining companies. Standard Oil Co. (Ohio) said it plans to spend $7 billion on Kennecott during the next 10 years for modernization and expansion. And Atlantic Richfield Co. said it will spend $500 million on Anaconda Copper Co. during the next five years. Another positive note was sounded by T. H. Adams of the United Banks of Colorado Inc. He said he expects the mining industry to experience moderate growth in 1982, led by "interest rate sensitive" demand from energy, electronics, and defense markets. And mining will be helped by "weak, but improved" construction and auto markets. The mining industry will be operating in an economy stronger than 1981, but still relatively sluggish by postwar standards, Adams predicted. Iron Ore Estimated US iron ore production in 1981 was 75.2 Mt (74 million long tons), up 6% from 70.7 Mt (69.6 million long tons) in 1980, though down from the 87.1 Mt (85.7 million long tons) produced in 1979. The mine value of usable iron produced from domestic mines was estimated at $3 billion. US iron ore was produced by 28 companies operating 35 mines, 26 concentration plants, and 16 pelletizing plants. The mines included 34 open pits and one underground mine. Byproduct ore recovered from copper- and titanium-mining operations accounted for less than 1% of total iron ore production. Fifteen mines operated by nine companies accounted for 94% of total

Jan 5, 1982

By A. Supomo, A. Partanen, M. Mular

Although many SAG operators may agree that on-line measurement of SAG feed-size is an attractive notion, applying this information in a control strategy that recovers the investment, engineering time, and maintenance is not necessarily straightforward. The P. T. Freeport Split-Online system was justified based on the assumption that SAG feed-size could be used in a feed-forward control strategy either at the regulatory or expert system control level. Something like “as the feed size changes adjust the throughput to avoid under or overloading the mill” was envisioned. However, following installation and calibration, Cluster Analysis showed that such an approach would not be successful. However, the sensor has been successfully integrated into Freeport’s grinding control strategy other ways. It provides an input to the Kalman filter parameter estimator used in a model-based control strategy that optimizes throughput and grind. The Split sensor also provides a characteristic feed size parameter for use in an on-line work index estimator. This paper describes the calibration and installation of the Split SAG feed-size analyzer, and reviews the evolution of its use in the Freeport C3 grinding circuit control strategy.

Jan 1, 2001

By W. F. McDermott

One of the major improvements in ore grinding economics in North America has been the introduction of autogenous grinding. The history of this has been well documented in previous AIME presentations and in the literature. Pickands Mather & Co. has been very active in the development of autogenous grinding since 1957. This early work led to the application of autogenous grinding to the Wabush Mines ore in 1965, with the installation of 6-24 ft. diameter by 8 ft. long - 1,750 HP mills. Development work recently completed in 1971 on another ore has led to the decision to install 5-36 ft. diameter by 15 ft. long - 12,000 HP mills when this project goes ahead. Between these two decisions lies a tremendous amount of research, development, and application of autogenous grinding to a variety of ores. The Pickands Mather Metallurgical and Research Laboratory in Hibbing, Minnesota has tested, in its pilot plant facility, ores from 9 different deposits. Some of these deposits have as much difference within their own ore types as there is between the various deposits. Therefore, it is safe to say that over 20 different ore types have been tested and flowsheets developed for commercial applications. Of the 9 projects, 3 are in operation. These are Wabush Mines in Labrador, Canada, (1) Savage River Mines in Tasmania, (2) and The Griffith Mine in Red Lake, Ontario, Canada. (3) Pilot plant work has encompassed over 300 individual tests. These tests were carried out in mills that were 5-1/2, 8, and 12 ft. in diameter. Grinding power requirements for the primary autogenous milling has varied from 2 to 30 KWH/Lt of crude ore.

Jan 1, 1972

By J. E. Moyle

Introduction In October 1983, Rio Algom Ltd. decided to develop the East Kemptville tin deposit. This decision was the result of intensive geological and metallurgical evaluation work, as well as careful studies of environmental and marketing aspects. The property is located about 54 km (34 miles) northeast of Yarmouth, Nova Scotia, near the village of East Kemptville. Geology The geological setting is a granitic intrusive at the contact with metasediment. The tin mineralization is in the form of cassiterite (SnO2), which is about 80% tin. Sulfides are associated with the deposit containing silver, copper, and zinc, which are recoverable as byproducts. The mineralization is associated with quartz veins and greisen alteration. In the center of the deposit, the zones are so close together they have the appearance of a massive or porphyry type deposit. Ore Reserves The mine plan is based on 56 Mt (62 million st) at 0.165% tin. Up to the ninth year, mill feed will average about 0.2% and lower grade material will be stockpiled. In later years, the average grade will be 0.16% tin. A 0.08% cutoff grade is planned for the entire 17 years of production. East Kemptville is a relatively low grade tin deposit, and only a combination of favorable circumstances has made it economic. • The ore body is situated at, or very close to the surface, permitting mining by open-pit methods at a very low stripping ratio. • The tin occurs in relatively large crystals that allow for good recovery by gravity techniques. • It is ideally located in a well-settled area, with all essential infrastructural requirements already in place. • The positive and cooperative response from various governmental agencies involved are encouraging. Mine Plan The mine plan incorporates an initial pit, followed by three expansions to the ultimate pit. Mining will be at 14 kt/d (15,000 stpd) and will include 9 kt (10,000 st) of ore plus 5 kt (5,500 st) of waste and low grade. After five years, the mining rate will increase to 21 kt/d (27,500 stpd). It is planned to drill vertical 165-mm (6.5-in.) blast holes on a 6-x 6-m (20- x 20-ft) pattern. Bulk slurry explosives will be used to break the rock. Loading will be with hydraulic shovels into 70 t (77 st) haulage trucks. The trucks will deliver the ore to a 13 x 20 m (42 x 65 ft) gyratory crusher, where it will be crushed to 150 mm (6 in.). Secondary and tertiary cone crushers will be used to further crush the ore to 13 mm (0.5 in.). Process Design Tin extraction technology is practically unknown in North America, but Rio Tinto Zinc, which has a 53% ownership in Rio Algom, is involved in tin production in the UK. Rio Algom has used this experience and expertise in metallurgical evaluation work and in tin-related aspects of plant design. Mineralogical and metallurgical test work have been performed, both on laboratory and pilot plant scales. Mineralogical work was done largely in the UK, the laboratory test program was carried out at the Lakefield research facility in Ontario, and pilot plant work was conducted in Australia. This program has resulted in significant improvements in recovery and concentrate grades, over best projections that could be made before this work was carried out. Concentrating Process The concentrator is being designed to process about 9 kt/d (10,000 stpd). The first step is to reduce the size of the ore to -700 µm (-24 mesh) in a rod mill/ball mill grinding circuit. The ball mill will operate in closed circuit with high capacity banana-type screens. The screened product will be deslimed to 30 µm (560 mesh) with cyclones, then passed over the pri-

Jan 4, 1984

By J. R. Dillon

Today more than ever, there is an increased demand for higher plant production capacities and efficiencies to offset the increasing costs for labor and materials. Plants involved in processing large volumes of bulk materials are particularly interested in more efficient material handling systems to (1) reduce labor and maintenance costs, (2) improve blending of non-uniform raw material feed to processing facilities, (3) improve blending of finished bulk materials from storage. DragScraper systems are being used effectively for handling a variety of bulk materials in both indoor and outdoor storage. Materials being handled include ores, chemicals, coke, clay, and stone, to name a few. Handling capacities these machines can provide vary from a few tons to over 1500 tons per hour. The basic components of a DragScraper machine consist of a bottomless Crescent scraper, a two or three-drum DragScraper hoist assembly including gas, diesel, or electric power unit, operating cables, guide blocks, and either fixed or mobile head and tail structures. They are presently available in sizes through 15 cubic yards, for operating spans up to 1000 feet.

Jan 1, 1969

By P. R. Shivdasan, R. D. Hagni

Cathodoluminescence microscopy (CLM) is uniquely applicable to the study of carbonatite-related fluorspar ore deposits. The mineralogy of the carbonatites, sodic and potassic fenites, marbles, quartzites and fluorite ores are readily determined and characterized. Calcite, dolomite, apatite and fluorite were deposited dominantly under high-temperature conditions, and they exhibit strong cathodoluminescence (CL). At the Okorusu, Namibia, carbonatite complex and fluorspar mines, three types of carbonatites have been recognized by field mapping and laboratory CLM. The main carbonatite intrusion is fine grained and consists mainly of calcite. A later phase of pegmatitic carbonatite consists of calcite, magnetite, Fe-rich pyroxene, apatite and pyrochlore. Pyroxene carbonatites were recognized mainly through their study by CLM.

Jan 1, 2002

By David J. Leeds

Increased interest in the utilization of the underground environment for water conveyance tunnels, vehicular tunnels, room and pillar type storage chambers, and military installations has focused attention on the possible effects of seismic activity on the stability of these types of structures. Earthquake experience has generally been related to the seismic stability on the surface of the earth. To establish aseismic design criteria, we attempt to predict the seismic exposure of a given site, both in terms of the spectrum of motion to which the structure must react favorably, and the recurrence rate of that motion. Experience gained from studying the effects of the Plowshare underground nuclear tests on existing tunnels and mines, coupled with the experience of the behavior of tunnels and mines in earthquake regions, provides us with an insight to predict the possible effects of earthquakes on underground openings and to establish design criteria. The type of damage ranges from minor fracturing and movements along pre-existing joints or faults through deformation of openings to collapse. The methods used to establish the seismic design criteria for an underground opening require modification of surface criteria to account for the difference in both amplitude and spectra underground. An aseismic evaluation should be performed on all important projects in areas exposed to possible earthquake motion. To perform this evaluation, information is required pertaining to the geometry and properties of structure and ground. Dynamic as well as static properties are required. A prediction of performance can be derived by exposing a model of the site, expressed in terms of its properties and geometry, to a design earthquake.

Jan 1, 1997

By James N. White

Completion of the Hall Street Coal Transfer Terminal in St. Louis marks the beginning of a major shift in the movement of coal to the power generating stations of the Midwest and South Central regions of the United States. The location of the terminal is significant. The 10 million ton- per-year facility is the first high capacity rail-to-barge transfer terminal to be built on the western shore of the Mississippi. This terminal and others to follow will link the massive coal fields of the West with the high coal consuming markets adjacent to the 25,000-mile Inland Waterway System.

Jan 1, 1979

By O. A. Bascur

The problem of setting the proper control objectives and/or having several objectives to be met in the implementation of grinding control strategies can be resolved. The design and implementation of a distributed control system for safe operation of a grinding circuit while maximizing profit is presented. An optimal recovery through- put target for a desired grind is accomplished by advanced control strategies at the DCS level. A mill load constraint control technique provides stable operation with a substantial increase in throughput, while the desired grind index is maintained by an inferential grind control technique without using a particle size analyzer. The integration of crushing/grinding and flotation in the overall design was the key to the success of the project. This synergistic effect was obtained by implementing a computer plant management system in the concentrator. The overall/philosophy was to keep the implementation simple while using new technology creatively.

Jan 1, 1991

By A. Ardiant, C. E. F. Mumbunan, I. Irawan

As a part of mine operation’s success key, Community Development nowadays seems to be a rule of thumb. Not only is for a short term need of the company during their operation but also for the national public opinion on the presence of the mining industry to a certain region. Nanggung region which is only a 2-hour drive from Jakarta, brings the challenges to Antam’s gold mine business in the areas such as illegal miner presence, easy access to the mine area as it is very close to the community, and national political sensitivity considering its proximity to Jakarta. The company allocates about 3%of its total operational cost annually for Community Development pro-gram. The program is derived through SWOT analysis and broad involvement of all community components to assure that the allocated fund is addressed to the right place. In the year of 2002, the company has successfully established a conducive-relation with the community.

Jan 1, 2004

By W. L. Jacobsen, T. L. Dyer

Cortez Gold Mines completed a feasibility study for a new openpit gold mine in north central Nevada. The production schedule calls for a peak in material movement. This schedule is based on current operational efficiencies and equipment parameters observed at the company’s existing openpit operation. To mitigate risks of over/under predicting production, simulation modeling was used to validate the loading and haulage requirements during the peak production period. The advantage of simulation is the ability to provide a realistic model of a mining operation based on scheduled and unscheduled random events during the period under analysis.

Jan 1, 2007

By K. G. Black

When one's attention moves from the control of a process unit or from the control of a department of similar units, to the control of an entire plant, two things must change. One is the techniques of control, and the other, the goal or purpose of control. The goal is no longer maintaining a certain quality of product or maintaining a particular flow rate through a process unit. The goal becomes economic - how to make the most money out of the plant. The principal control techniques become those of optimizing, rather than regulating. It is quite common to discuss the control of an entire plant in terms of three levels of control l; regulation, supervision, and optimization. Regulation is the control of inputs to a process unit at desired values. This is accomplished by two or three mode analog controllers or DDC loops. The second level of control is the selection of these desired values, e.g., set-points of the regulation controllers. The third level is to make such adjustments or corrections to the supervisory calculations so as to optimize the plant profit. It is immediately evident that supervisory control and optimizing control are intimately related, and a distinct dividing line cannot always be drawn between the two. Some authors prefer to add even more levels such as adaptive control and management information systems, but we believe our point is well made without further division of the functions which are necessary for complete plant control. To accomplish the economic goal, that is, the optimizing (maximizing) of the profit of a mineral processing plant, the plant should ideally be operated as if it were a separate company, buying its raw materials on the open market and selling its products on the open market. While petrochemical plants endeavor to operate in that manner, unfortunately this is not true of a mineral processing plant. The plant manager has to accept the amount and quality of ore the mining department delivers to him, and the products (concentrates) are delivered to another department of the company for further refining. In some cases he may sell

Jan 1, 1972