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By A. D. Davis

A ground-water flow model was developed to simulate the post-mining recharge rate at the Belle Ayr surface coal mine in the Powder River Basin near Gillette, Wyoming. State regulations of Wyoming require that reclaimed lands should provide a rate of recharge that approximates the pre-mining rate. Steady-state calibration and transient simulations showed that the post-mining recharge rate in backfill is at least five centimeters per year, exceeding the original infiltration rate of 0.4 centimeters per year to shallow, overburden aquifers. Major reasons for the increase include the flat topography of reclaimed areas and decreased spoil density. It is expected that the backfill will disaggregate and settle over time, eventually resulting in a nearly constant recharge rate less than five centimeters per year but somewhat greater than the pre-mining rate.

Jan 1, 1995

By N. Rajmohan, K. T. Jacob

Presented is a thermodynamic feasibility analysis of extracting base metal chlorides from low-grade, multi-metallic oxide ores using CaC[1]2 as a chlorinating agent in the presence of SO2 and O2. The oxides react to form corresponding chlorides, while CaC[1]2 is converted to CaSO4. The Ellingham diagram is used for comparing the standard Gibbs' free energy change for the sulfation-chlorination reaction of a large number of oxides. Except for alumina, silica and chromia, most of the other metal oxides will be converted to their respective chlorides. The volatile chlorides can be condensed, and the chlorides present in the condensed state can be leached. A process is proposed that uses a nontoxic chlorinating agent and gives an efficient separation of the metallic values from the gangue.

Jan 1, 1993

By P. D. Rao, D. E. Walsh, W. W. McClintock

A column flotation system was designed to conduct release analyses on a very finely disseminated, stratiform copper ore from the Denali Copper Prospect in Alaska. Chalcopyrite is finely disseminated in a carbonate matrix, and pyrite is disseminated in both the chalcopyrite and the gangue. Grinding to -20 µm is required for adequate liberation. A 25-mm x 2.4-m glass flotation column was designed. The ore and its flotation tailings were repeatedly floated to obtain a rougher concentrate. The rougher concentrate was then repeatedly floated to produce release fractions. These fractions were analyzed for copper, iron, calcium and magnesium, and release curves were plotted. The procedure was replicated at each of two grinds, i.e., 99% and 88% passing 37 µm (400 mesh). The procedure should find application for flotation characterization and process selection. Bond grindability testing was also conducted.

Jan 1, 1996

By Wen L. Cai

This paper examines the effects of different block dimensions and cone bases on the economic open pit limit design based on case studies on a coal deposit consisted of two west- east elongated, near vertical thin seams starting with a base model dimension of 200x170x50 (NX*NY*NZ) with a block dimension of 10mx5mx4m (DX*DY*DZ), the blocks were combined along easting, northing and elevation directions. Two cone bases were used, one with a base radius of 10 meters, the other with a base radius of 0. Two pit design algorithms, one is the floating cone algorithm, the other is an optimum algorithm, were used to see which one of the two pit design algorithms is more sensitive to block dimension and cone base variations. The results are presented and analyzed.

Jan 1, 1992

By J. West-Hansen

The problem of long-term production scheduling in an underground coal mine is approached through the concepts of sequencing theory. Based on a typical layout of mining sections, the principles for the approach are: 1. The definition of mining sections as jobs, and mining methods as machines. 2. The reduction of the overall mining activity to the general problem of nonpreemptive scheduling of precedence related jobs on parallel machines. 3. The simplification of the layout to a general type directed network with the nodes representing jobs and the arcs representing precedence relations. 4. The determination of a feasible and near-optimal job schedule given the objective of maximizing the total net present value of the jobs. A general scheduling heuristic is being developed for implementation on an IBM personal computer (PC) as part of an interactive computer system for long-term production planning.

Jan 1, 1986

By J. Hanna, I. J. Anazia

The Mineral Resources Institute (MRI) has developed a unique process for selective fatty-acid flotation of phosphate minerals from siliceous ores. The process takes advantage of preferential adsorption (over silica surfaces) and high reactivity of fatty-acid collectors toward apatite under slightly acidic conditions (pH 5 - 6.5). The process was successful in upgrading various Florida and North Carolina phosphate samples containing 7.7% to 15.3% P205 and 62.3% to 77.2% siliceous gangue (Insol). The resulting concentrates assayed 28.1 % to 33.1 % P205 and 4.1 % to 9.9% Insol, with high P205 recoveries of 81.8% to 92.2%. Depending on the degree of surface contamination of the silica particles in the flotation feed, one or more cleaning steps, with or without sodium silicate, would be required to achieve higher-grade phosphate concentrates.

Jan 1, 1994

By F. M. Doyle, Jun Li, Milton E. Wadsworth, E. Peters

The Raman spectroscopic technique was used to study electrochernically oxidized pyrite in-situ. It was found that Raman-active sulfur oxidation products could be detected after a proper potential was applied to the pyrite electrode for a threshold period of time. There existed a minimum potential below which sulfur could not be observed. However, when the applied potential exceeded a certain limit, the intensity of the sulfur spectra decreased. The minimum potential at which sulfur could be observed changed with solution pH. For both acid (pH 2.7) and weak alkaline (pH 9.2) solutions, the minimum potential was 0.8 V/SCE, whereas for neutral (pH 6.5) solutions the value was 0.4 V/SCE. In strong alkaline solutions (pH 12.5) surface sulfur oxidation products could not be detected at any potential.

Jan 1, 1993

By H. Y. Yang, L. L. Tong

Semicontinuous-flow, multistage biooxidation experiments of a Chinese refractory, high-sulfur gold concentrate were conducted with different pulp densities for gold recovery. The maximum iron, sulfur and arsenic oxidation rates of 90.0, 88.8 and 96.4 percent, respectively, were obtained in the experiments with 5 percent pulp density. The highest gold and silver recoveries of 96.1 and 68.9 percent, respectively, were achieved by cyanidation. The negative effect of high pulp density on biooxidation was related to the decrease of free and adsorbed cell concentrations with increasing pulp density. The data of iron oxidation from the semicontinuous biooxidation experiments were best fitted by a shrinking core model in which the chemical reaction controlled the rate. The dissolution kinetics of sulfur and arsenic were therefore analyzed by a shrinking core model in which the internal diffusion controlled the bioleaching rates. For the quantitative description of biooxidation efficiency, two parameters — mass dissolution rate and volume dissolution rate — were proposed. In the first stage, the mass dissolution rates of minerals were negatively correlated while the volume dissolution rates were positively correlated with the actual pulp density. The iron and sulfur mass dissolution rates decreased with increasing actual pulp densities in the third and fourth stage, while linear increasing relationships between the volume dissolution rates of minerals and the actual pulp density were observed from the second to fourth stage. The highest mass dissolution rates of minerals were obtained in the first stage with 5 percent pulp density. The highest volume dissolution rates of minerals were obtained in the first stage with 15 percent pulp density. Compared to the volume dissolution rate, the mass dissolution rate was more intuitive to reflect the negative effect of high pulp density on biooxidation efficiency.

By Enrique Farjeat P.

Mexico City's growth has encourage development of underworks to meet drainage and transport requirements. Deep Drainage System has been constructed through a complex tunnel network, 108 km in total, varying from 30 to 12 m in depth and 6.24 or 4.00 m in diameter. Wide range of stability conditions presented in soils has forced the use of several tunnelling methods, going from the manual procedure by the use of pneumatic tools, to slurry shields, or open face shields, with the aid of or without compressed air. For transport, subway tunnels have been excavated in very compact silty sands, 9.15 and 9.51 m in diameter, with the use of open shields one of which was designed by Mexican engineers as a prototype machine. This paper tries to describe evolution of tunnelling methods used in Mexico City during last 20 years.

Jan 1, 1991

By Yan. T. Y.

Uranium precipitation from an eluate using hydrogen peroxide has been studied in a laboratory continuous unit. The important process parameters were the dosage of H2O2, reaction pH and chloride ion concentration. In the process, the alkaline eluate is acidified to a pH of 2, then H2O2 at 110%-130% of the stoichiometric requirement is added, and finally NaOH is added to maintain a pH of 4 for one hour to complete the precipitation and settling. The chloride content in the system is controlled to 50g/L or below by rejecting part of the decant solution. The process has been successfully employed in commercial operations. The uranium recovery is complete, while the precipitate settles rapidly and can be processed to high purity yellow cake without filtration.

Jan 1, 1991

By H. Nemutolluhi

Discussion by G.E. Agar In the April issue of Mining Engineering there is a technical note "New size laboratory ball mill for Bond Work Index determination" written by M. Nematollahi which is very confusing. My recollection is that a description of an experimental procedure should be such that someone skilled in the art should be able to reproduce the results. Applying this test to this particular article would according to me cause a failure to be registered. There are apparently three batch laboratory test procedures referred to in the article. One of these is the traditional Bond test with all its limitations such as dry grinding followed by exclusively dry screening, two attributes which fail to simulate conventional industrial grinding. There is a second procedure (NSBM) which the author is advocating as an improvement of the Bond procedure but from the description I doubt that anyone would be able to repeat the procedure. For instance, is the grinding to be done wet or dry, if wet at what pulp density, and how is the screening to he done, wet or dry or a combination? On top of that uncertainty the author refers to a simplified "Anaconda" test. No reference is provided for this procedure and in the text the results of the technique are referred to "Denver wet ball mill" results and say nothing of Anaconda results. Reply by H. Nematollahi I think that the method proposed to determine the Bond Work Index in this paper is clearly described in the paper as follows: •The method of grinding is dry (Ref. paragraph I, subject Method description); •The size of ball mill, the quantity and size of balls are noted (Ref. paragraph 1, subject Test equipment); and The test procedure is described in detail (Ref. paragraphs 1, 2 and 3, subject Test procedure). As it is noted in the paper, the test has been worked out under the same condition as Standard Bond Test with ball mill, so dry (screening should be dry also), but using a smaller ball mill and a smaller quantity of sample. Considering the above mentioned points, I believe that the method has been fully and clearly described and it ought to be repeatable as well. The Anaconda test is a method for determination of Work Index and could be done by any types of laboratory wet ball mill (Yan, R.F., et al., cf. References). I applied this method using a laboratory Denver ball mill (Ref. paragraph I subject Determining the Bond Work Index). So the results referred to Denver wet ball mill can be equated the same as Anaconda method.

Jan 1, 1996

By R. R. Fernández

Newmont Mining Corporation has been processing refractory gold ores by roasting since January 1995. The ores are refractory due to the association of the gold with iron sulfides, namely pyrite and marcasite, and the presence of carbonaceous components and is oxidized in an oxygen-enriched gaseous atmosphere. The structural characteristics of the calcine have a strong influence on the eventual recovery of gold by cyanidation. These characteristics are strongly dependent upon the roasting conditions, mainly the roasting temperature. Originally, the sulfides and carbonaceous matter were oxidized at a temperature that typically ranged between 535 and 560 °C. Controlling the roasting temperature within the lower and narrower range of 510 to 515 °C produced a calcine with a larger surface area(e.g. higher porosity), which resulted in a gold recovery increase of3.3 percent in 2001 equivalent to 5.1 million dollars of additional gross revenue. The new strategies of operating at temperatures never tested before are discussed. It constituted a great challenge to the control room operators for the lower operating temperature falls much closer to the ignition temperature of the sulfide ore and smaller and more frequent adjustments in operating parameters were needed to improve and maintain high gold recoveries. Based on the encouraging results that were achieved in 2001, a decision was made to install a “smart” process control system because the roaster temperature control was performed by manual inputs from the control room operators. The development and implementation of the control system to reduce or eliminate the temperature variability between the different control room operators commenced in June of 2002 and was fully operational by the end of October. Lower roaster temperatures were achieved (500 to 505 °C), which resulted in further increase of gold recovery by about two percent. The overall recovery increase equated to 8.0 million dollars of additional revenue for the first 10 months of 2002.

Jan 1, 2003

By Roger Holmberg

INTRODUCTION The driving of drifts is a very important aspect of underground mining. It is not unusual for the percent¬age of rock broken during development in a mine using sublevel caving, for example, to be as much as 25% of the total. If one also considers the amount broken for transport, ventilation, and exploration drifts, one can easily understand that the planning and excavation of drifts play a major part in the total economics of the mine (Fig. 1). Increasing mechanization in mining demands larger tunnel areas for transport and mining equipment. With modern machines the hard work involved in using hand¬held pushers is gone, and a better environment is achieved. More rational methods could be used, but much of the experience the working man acquired by working close to the rock face (such as utilizing the natural weak planes in the rock when he placed the drill¬ing holes) has unfortunately been lost. By having sepa¬rate shifts for drilling, loading, and hauling, more atten¬tion has to be placed upon a well-designed drilling pattern. Some reduction in the number of holes required can be achieved with mechanized drilling because of the larger holes that can be produced. On the other hand, it is probably not possible to achieve the same precision as with pneumatic pushers, and it is difficult to utilize the larger holes because these cause more damage to the re¬maining rock. Recently, however, the precision has be¬come very good with the parallel booms and automatic devices for setting the lookout angle (Fig. 2). A larger arch of the drift roof requires a more carefully executed blasting procedure than before in order to prevent rock fall and to insure a sufficiently long stand-up time. In this chapter, empirical relationships that can be used to design an economic and optimal drift blasting design will be presented. The principles of the calcula¬tion method are based upon the earlier work of Langefors and Kihlström (1963) and Gustafsson (1973). COMPARISON OF EXPLOSIVES To provide for the use of various explosives it is necessary to have a basis of comparison. Several meth¬ods have been developed to characterize the strength of an explosive. Some examples are comparison of values given by (1) calculated explosion energies; (2) the ballistic mortar test; (3) the Trauzl lead block test; (4) the brisance test; (5) the weight strength concept; and (6) the underwater test. However, most of these meth¬ods should be used carefully when stating the breaking

Jan 1, 1982

By Robert L. Frantz

The various facets of university-company cooperative activities and the opportunities for collaborative research for the mineral industries are presented in this paper. The advantages and disadvantages of various alternative research funding arrangements for the researcher, the university, and the sponsoring corporation are discussed. Examples of university-industry cooperative research programs are briefly outlined. The concept of a cooperative mining research program is presented. The improvement of performance through techno- logical developments and human skills may well be the greatest competitive edge available to the U.S. mining industry. In this regard, the need for new endeavors between industry and universities can hardly be overemphasized.

Jan 1, 1984

By D. S. Kostick

Soda ash production in the United States during 1994 increased 4%, reaching 9.3 Mt (10.2 million st). That was the second highest production year. Domestic nameplate soda ash capacity is about 11 Mt (12 million st). Overall, the industry operated at about 85% of capacity. Some companies produced at 100% of capacity while others produced at below their design ratings. Although domestic soda ash consumption was slow during the first half of the year, third- and fourth-quarter sales increased, especially in the flat¬glass and fiberglass sectors. These two sectors increased because of greater demand for glass products in the automotive and construction industries. Soda ash sold for pulp caustic soda increased during 1994. This was due to higher prices for and reduced availability of electrolytic caustic soda. It is often preferred for pulp digestion because of its stronger alkalinity.

Jan 1, 1995

By William L. Barnes, Maurice Leiser

INTRODUCTION The literature on project management considerations and techniques for large rail rapid transit projects in mid-1979 was fairly large and getting larger. UMTA was establishing special organizational elements to deal with the subject as it related to those properties requesting capital assistance, and comprehensive project management plans were being requested from those properties as an early step to build confidence in UMTA that the particular project could be planned, designed, and executed efficiently and effectively, thereby using critically short capital assistance funds to best advantage. In the late 60's and early 70's, the choices of models for heavy rail construction would have been quite limited.., and perhaps much simpler than was the case in 1979. One of the writers was with an agency in 1979 in the process of finalizing its organization for project implementation after several years of study and had the opportunity to review the experiences of others at length, presumably to build on the success and to try to avoid the pitfalls and failures. The properties that were surveyed all reported more or less success stories. One could only conclude that there are many ways to organize for a major transit (construction) project and that there is a great variety of viable management structures, but it seemed that chances for success were best where the executing property had prepared a comprehensive project management plan, then resolved to follow it. Perhaps more important, a commitment or prior approval had been sought from UMTA as to the validity of the project management plan, and the property was allowed to proceed without detailed prior UMTA review or approval of all specific intermediate steps. Most important, however, an early overall UMTA funding commitment had been obtained that permitted some reasonable expectation of the rate of funding, perhaps by commitment

Jan 1, 1981

By A. C. Nixon, A. D. S. Gillies, Hsin Wei Wu

Ventilation and refrigeration strategies for the 3000/3500 orebodies, Mount Isa Mines, Australia were reviewed. The orebodies lie between 1,200 and 1,800 m below surface and are characterised by steep geothermal gradients, intensively mechanised stoping methods and high ambient air temperatures in summer. Providing acceptable working conditions under these situations is difficult and costly. The ventilation air quantity required in the 3000/3500 orebodies exceeds 550 m3/s when the production rate is greater than 2.5 million tonnes per year. The maximum quantity that can be supplied and exhausted using the existing intake and return ventilation system is estimated at about 630 m3/s. For production rates greater than 3 million tonnes per year, either new intake and return airways will need to be developed or innovative approaches such as controlled partial recirculation used. A description is given of technical and economic investigations conducted at Mount Isa into controlled partial recirculation. During these investigations, a major field trial of a controlled partial recirculation system was undertaken with results indicating that an acceptable working environment can be achieved practically and economically utilising this approach. The cost involved in controlled partial recirculation for supplying the required ventilation quantity when the production of 3000/3500 orebodies exceeds 3 million tonnes per year is examined. A cost comparison between recirculation and other conventional alternatives such as new surface shaft or underground booster fans is included. The results indicate that the option of a new surface shaft will incur a high initial capital cost and moderate annual operating costs. The alternative of underground booster fans will necessitate a moderate capital cost and very high operating costs. Controlled partial recirculation, in this case, offers both the lowest capital and operating costs of the three options evaluated.

Jan 1, 1995

By Oleg N. Tikhonov

A new extraction method which includes a combination of gravity and alternating electromagnetic fields with interfriction of mineral particles was proposed (Tikhonov O.N., 1979) and tested in lab conditions at Mineral Processing Department of St. Petersburg Mining Institute. At the beginning of the work the main goal was to find an efficient alternative to traditional extraction methods (like shaking table, etc.) used for concentration of fine alluvial gold particles. After the method and a series of gradually improved lab separators (Tikhonov, et al., 1983) were positively tested for gold (Tikhonov, et al.. 1986) the efficiency of the method was also proved for extraction of and other heavy minerals from alluvial and ground ores in experimental work at the St. Petersburg Mining Institute. The fine heavy minerals were separated to concentrates with high technological indices (Andreev et al., 1987), (Tikhonov, Gladkov, 1987), (Tikhonov, Gladkov 1989). In parallel a semi-industrial gravi-electro- magnetic separator was made and several technological flowsheets were designed for testing the method in industrial conditions (Tikhonov O.N., Miazin B.N., 1987), (Tikhonov O.N., Tolstunov S.A., Jebelev V.V.,1989). A series of industrial tests accompanied with complete sampling and standard technological measurements were performed at several alluvial deposits, with good technological indices. Thus the method and the flowsheets displayed high additional recovery of fine alluvial gold particles during elkrated industrial experiments in Siberia (Aldan province). At higher concentrate grade this method, even in one-stage operation, increased gold recovery more than by 25% compared with shaking table (for particles -0.10 mm gold recovery was 82% vs. 55% by shaking table, for particles - 0.25mm the recovery was 97.5% vs. 61.8%). Experiments were performed by joint team of St. Petersburg Mining Institute and the Irkutsk Institute of Rare Metals (Tikhonov O.N., Gladkov S.A., Tarasova T.B., 1991).

Jan 1, 1995

By B. Pollack, L. G. Stevens

INTRODUCTION In other chapters of this volume many of the important technical aspects in the design of concentration and dewatering circuits have been discussed. This chapter will deal with the steps necessary to bring that information together in order to produce the specifications necessary to purchase the equipment and complete the design. It deals with the plant as a whole rather than just concentration and dewatering, since the steps are essentially the same. Prerequisites to Engineering Design Feasibility studies will have been completed and design parameters established prior to the start of the engineering design stages for a metallurgical plant. The data base for the feasibility studies will have included laboratory and/or pilot plant test work, historical data, proposed plant site data such as climatology, topographic mapping, geological and geotechnical information. These studies will have established the general project design criteria as well as the process design criteria leading to the preparation of preliminary flowsheet ( s) and/or alternative process flowsheets. These will have incorporated metallurgical and material balances, allowing identification of major equipment items, and development of preliminary plant site plot plans. Preliminary technical and economic assessments and recommendations will have been prepared during this phase for Owner's review and evaluation of the technical, physical, and cost effectiveness of the overall project and its component process, process facilities, and supporting facilities elements. After Ownert s approval and decision to proceed, the engineering design can be initiated, using the above information as the basis. Engineering Design Phase The engineering design of metallurgical plants and their elements are generally developed in two sequential stages, namely: (1) Basic ("Front end") Engineering Design (2) Detail Engineering Design Figure 1 relates these engineering stages to the entire program from feasibility through construction.

Jan 1, 1986

By Joseph E. Worthington

The Haile Gold Mine is one of the earliest gold discoveries in the eastern United States as well as one of the largest gold producers from that region. Gold was first discovered in the gravels of what is now known as Haile Gold Mine Branch in 1828 on lands then owned by Colonel Benjamin Hale (spelling not the same). Small placer workings were immediately started and it was not long before the surrounding residual soils and saprolites were also found to contain gold. Following the presence of gold in the residuum eventually lead to the discovery of the underlying gold-bearing lodes that actually outcropped in saprolite form where not masked by heavy vegetation or a thin veneer of Coastal Plain sand. The earliest attempts at production, immediately after discovery, utilized slave labor, washing gravel and residuum from leased tracts that were 15m (50 ft) square. Several at- tempts at small milling operations were made during the en- suing decades, but there was really no substantial or successful production until the 1880s.

Jan 1, 1990