Search Documents

By Peter G. Chamberlain

Many mining engineers considering in-place leaching for the first time are uncertain as to how to evaluate ore bodies for potential leachability. Evidence presented in this report emphasizes the critical role of permeability in such evaluations. Whether an engineer is considering leaching "tight" formations typified by the porphyry copper deposits of the southwest or the permeable uraniferous sandstones of Texas and Wyoming, permeability measurements are the keystone of any testing program. Permeability tests range from cheap and easy to highly expensive and sophisticated operations. Constant head, variable head, and pumping tests all have application in evaluating ore bodies for in-place leaching. Costs of these tests will not unduly deflate the pocket- books of most mining companies. The basic procedures for selecting and running these field tests-including preparing suitable wells--are reviewed for the benefit of newcomers to in-place leaching.

Jan 1, 1979

By Harry R. Dahlberg

Heap leaching is discussed as a unit process. Five factors are identified as the principal elements that can be measured and/or controlled. These are: 1. Material Size 2. Volume of solution 3. Tonnage of heap 4. Time 5. Solution concentretion. The Rate of Extraction Curves obtained from an actual heap or column tests describes the economic viability of the unit process. By analysis of these curves and the five factors end their inter-relationship it is possible to determine how the extraction rate can be increased. The practice of leach and rest used by many operators is analyzed and shown on a hypothetical rate of extraction curve. The influence of the driving force of the hydrogen ion and ferric Lon as Lixivants and its effect on the rate of extraction curve is discussed.

Jan 1, 1979

By Max L. Glenn

As part of its goal to minimize requirements for mineral commodities through conservation and substitution, the Bureau of Mines conducted research to develop molybdenum as a potential substitute for superalloys containing the strategic and critical materials, nickel, chromium, and cobalt. Since molybdenum requires protection from oxidation at high temperatures, several complex oxides of the spinel and perovskite structure were investigated to determine which possessed the greatest potential for use as oxidation-barrier coatings. Two spinels (MgCr204 and MgA1204) and five perovskites (BaHfO3, YCrO3, NdCrO3, BaZrO3, and LaCrO3) were selected for testing. Various high-temperature prop¬erties that are important for oxidation-resistant coatings were examined, including diffusion rates of oxygen and molybdenum, recession (vaporization) rates, hardness, erosion resistance, sinterability, reactivity with molybdenum, and ease of application. From the evaluation of high-temperature properties, two spinels (MgCr204 and MgA1204) stand above the other oxides as possible coatings for molybdenum. However, laboratory coatings of MgCr204 at 1,360° C (2,480° F) are equivalent or inferior to currently available MoSi2 coatings; laboratory coatings of MgA1204 may be comparable under limited conditions to currently available MoSi2 coatings at the same temperature.

Jan 1, 1979

By William J. Karwoski

To improve rock-burst control in deep vein mining, a field demonstration has been carried out by the Bureau of Mines to test a new rock-burst control concept--rock preconditioning in advance of mining. The procedure is to fracture the major rock-burst-prone zone during the development of each level before production mining. This allows initial mining to begin in a preconditioned zone and later mining from the level below to end in a prefractured or preconditioned pillar zone that is relatively burst-free. This full-scale field demonstration was carried out On the 7700 level of the Hecla Mining Co.'s Star-Morning mine. An expected burst-prone zone along the main vein between 7 and 10 crosscuts, some 250 feet long by 40 feet above and below the level, was preconditioned by blasting a fanlike ring of drilled holes from each crosscut. The drill holes, up to 100 feet long and 4 inches in diameter, were loaded with high explosives and shot. About 4,550 feet of drilling and 9,000 pounds of explosives were involved in the effort. Comparison of preshot and postshot data from monitoring instrumentation indicated that the blasting did fracture the rock in the planned zone and did result in an immediate stress decrease and softening of rock. The field data compared favorably with predicted results from a finite-element computer model simulation of the destressing. Based on observations made during mining and on the data from the Hecla rock-burst monitoring system, there is no doubt that this preconditioning in advance of mining demonstration resulted in much-improved rock-burst control during first-floor mining. It was also concluded that preconditioning the vein was more effective than preconditioning the wall rock.

Jan 1, 1979

This section is addressed to the computing needs of corporate (top) management which are less specialized than the computing needs of operating or divisional management described in the previous sections. It is essential that the computer be used as a tool to serve corporate management and that information and results are presented in a succinct form. A key responsibility of the manager of computer applications is to communicate well with corporate management. A common trap for computer management is to allow the development and implementation of sophisticated and academically interesting computer applications which do not enlighten corporate management in its decision-making task. Thus it is essential to devote sufficient time and effort to designing and specifying a computer system before developing and implementing the system.

Jan 1, 1979

By J. R. Riding

In situ solution mining of uranium has environ- mental advantages over conventional mining. The leaching of uranium, however, alters the groundwater quality in the aquifer where the mining occurs. Currently, regulations require that the groundwater be restored to approximately its original condition. This paper, a summary of a study funded by the Bureau of Mines, reviews the state of the art in restoring groundwater quality after in situ uranium leaching. Current restoration practices discussed include disposal of liquid wastes in deep disposal wells and evaporation ponds, producing from all wells during restoration, and recirculating water purified in surface plants. Methods for predicting the effectiveness and cost of current techniques are presented. Possible alternative techniques are also described. Two restoration operations are discussed.

Jan 1, 1979

By Ronald L. Wiegel

In the 2000 page Handbook of Mineral Dressing by A. F. Taggart, the subject of material or metallurgical balances is covered in only six pages of eve-straining type. Although Taggart's work was compiled before 1945, and there has been significant progress since then in most areas of mineral processing, the practicing metallurgist's approach to metallurgical balances remains exactly the same as described in those six pages. There are several portions of Taggart's text appropriate to introduce the subject of this paper which I would like to quote: "Milling Calculations. Computations of per¬formance in milling are complicated by the fact that it is a continuous rather than a batch operation, that the quantities of solid materials handled are large and usually mixed with water, hence difficult or impos¬sible to weigh. Fortunately it is possible to determine many facts concerning perfor¬mance without knowing weights, if the value of various constituents is known in some common unit, e.g., the content of some par¬ticular metal or mineral, or of water, or some ingredient soluble therein, or of par¬ticles of a particular size or falling within some particular size range, or the like."

Jan 1, 1979

By Leons Kovisars

Two-thirds of the world copper produced in 1979 incurred cash costs between $0.35 and $0.89 per lb, averaging $0.62 per lb. North American costs are higher at $0.48, $0.88, and $0.68 per lb, respectively. Variable cash costs are $0.20, $0.58, and $0.39 per lb for the world; $0.29, $0.63, and $0.46 per lb for North America. Cash costs are relevant for comparing existing operations and evaluating market pricing during periods of excess supply. Analysis of full costs of production (cash costs plus income taxes, profit, and non-cash items) becomes relevant in evaluating the incentive price required to develop new production plants. In mid-1979 this cost (incentive price) was about $1.10 per lb; it is expected to approach $2.00 per lb by 1985.

Jan 1, 1979

By Marvin Barnes

The Sacaton mine is a relatively new open¬pit copper mine owned by Asarco, Inc., which has been in production only a few years. The mineral deposits consist of two ore zones; a western, near-surface zone minable by open-pit methods, and a deeper eastern zone which will have to be mined underground, probably by block caving. The copper deposits are related to small quartz monzonite intrusives implaced into Precambrian granites. Both rock types act as hosts for the copper mineralization, which has been secondarily enriched sufficiently to constitute ore. In the western zone the enriched supergene chalcocite has been subjected to postenrichment oxidation processes with partial destruction of the chalcocite blanket. The deeper eastern chalcocite zone has not been as severely affected by later supergene processes. Several mineralogical populations are represented in the total copper deposit. Leached capping overlies the chalcocite enriched zone, and in the western ore body there is a very irregular interface between the leached and enriched zone, which is complicated by oxide copper minerals irregularly distributed. Primary lower-grade chalcopyrite mineralization occurs below the enriched zone. When mining commenced near the top of the enriched zone, considerable difficulty was encountered in correlating the actual grade of ore in place with the estimated grade of the mining blocks. The mineralogical complexity and fairly wide spacing of diamond drilling made reliable grade estimation of individual ore blocks a difficult procedure.

Jan 1, 1979

By Bruce T. Stanley

Though most porphyry-molybdenum deposits lend themselves to various reserve estimating techniques, still there must exist the constant effort to not merely obtain an estimate but rather to obtain the best estimate. To answer the challenge of a complete ore reserve estimation, one must attack the problem with the purpose of obtaining the results closest to reality. What we wish to determine is simple: metric tons (short tons), grade, and kilograms (pounds) of molybdenum contained. In order to effectively accomplish this task, further consideration must be given to the ease of using the information obtained and its inherent reliability. The only known method for ore reserve estimation that supplies the needed results along with its distribution and a measure of confidence provided by sound mathematical backup is geostatistics. The following review is an example of geostatistical application of an ore reserve estimate at the AMAX Henderson mine demonstrating technique, viability, and usefulness of results.

Jan 1, 1979

By W. H. Gross

To pollute means to make impure or unclean. We tend to forget that man defecates, urinates, sweats, expectorates and exhales and in himself is a beastly little pollutor-but then so is Mother Nature! The object of this paper is to try to put what man and nature do into perspective. The facts indicate that although man may sweat a little, Ms. Nature is by for the greatest pollutor of all.

Jan 1, 1979

By Lamont G

The significance of blasthole stemming in determining blasting efficiency has been extensively studied for small diameter holes using gelignite explosives. Recent developments in blasting technology with the use of large diameter holes and low cost blasting agents have not included studies of the effectiveness of different types of stemming materials. A review of previous measurement techniques and the relevance of these earlier results to present stemming practice is presented in this paper. A survey of operating mines in Australia provided information for this analysis. The value of stemming in reducing blast noise and improving blasting efficiency is generally accepted but the use of bulk-loaded explosives in underground operations has resulted in the practice of eliminating the use of stemming material.In surface operations drill cuttings are used for stemming on the basis of availability rather than efficiency in providing resistance to movement. The development of laboratory testing techniques, based on noise level and borehole pressure measurements, to evaluate the relative efficiency of different types of stemming materials is described. The results of an extensive series of tests with ANFO explosive showed similar trends tothose obtained from earlier published results with gelignite type explosives. Conclusive quantitative results on the relative effectiveness of different stemming materials will require minor modifications to the existingequipment. Observations and the results of pressure measurements indicated the presence of shock wave effects which caused the ejection of the top section of the stemming column when using coarsely fragmented material.

Jan 1, 1979

By K. V. S. Sastry

Pelletization, sintering, and pressure compaction are the most common methods used to agglomerate ore concentrates. Over the last 20 years agglomeration has grown enormously in popularity and has become a standard unit operation in the nonferrous, cement, fertilizer, and ceramic industries. However, it is in the processing of iron ores and the preparation of blast furnace feed that agglomeration is used most and its application has grown fastest (Ball, et al., 1973; and Sastry, 1977). This chapter deals with an account of the current practices and future trends in the application of computers relative to an overview of the process design and engineering considerations of the unit operation of agglomeration.

Jan 1, 1979

By Staff

This Bureau of Mines report is devoted to descriptions of investigations pertaining to underground mine trailing cables, their handling, splicing, and the use of couplers.

Jan 1, 1979

By F. Brost

In 1976, a decision was made to install a conveying system to transport ore 10.3 km (6.4 miles) from the new Eisenhower Partnership Mine to Twin Buttes. The system, supplied and installed by Cable Belt Limited, of England, represents a radical departure from conventional conveyor systems. The driving tension is applied through two endless wire ropes while the belt acts only as the carrying medium. Features of the system are a single 2240 kw (3000 hp) drive and a 60° angle bend. Construction was completed in two years. Full production was attained after several months' operation.

Jan 1, 1979

By Peter E. Hart

The search for mineral deposits is a difficult high-risk enterprise whose likelihood of success is influenced by a great many factors. Among the most important of these factors are the knowledge and judgment brought to bear on an exploration problem. If, for example, a panel of expert highly experienced geologists could be convened to examine every prospect, it is at least plausible that fewer deposits would be overlooked in the exploration process. Unfortunately, whether a project is undertaken by a private company or a government agency, it is not usually feasible to convene such a panel. Our purpose in this chapter is to describe an experimental system of computer programs called PROSPECTOR that is intended ultimately to provide expert consultation on problems of mineral exploration and resource evaluation. Ideally, PROSPECTOR would serve as a surrogate for a panel of expert economic geologists. However, because the cost of computation is very low and is projected to drop by a factor of about 100 over the next two decades, systems like PROSPECTOR could be made available to the exploration enterprise at large. When consulting about a prospect, the PROSPECTOR system first gives the field geologist an opportunity to tell the program about the most significant features of the site: the major rock types, minerals, and alteration products. The system then compares these observations with a number of stored models of various types of ore deposits and forms alternative hypotheses about the types of deposits that might be present. The system next engages the geologist in a dialogue in order to obtain additional data that would be useful in resolving the multiple hypotheses. The geologist can interrupt the system at any time to obtain clarifications of questions, volunteer additional data, request the reason a particular question has been asked, or obtain a general summary of the state of the consultation process.

Jan 1, 1979

By James Wm. White

Calculation of consistent steady-state mass balances in operating metallurgical plants and pilot plants is of interest for many reasons. In general, enough data can - and should - be collected to overdefine the system, i.e., permit calculation of nonmeasured quantities via more than one path. In such cases, resulting steady-state metallurgical mass balances rarely close due to sampling, analysis and/ or instrument error. Indeed, determination of a valid calculational sequence for complicated process flowsheets with multiple recycle streams can prove quite difficult. Here we present and discuss a method that has proven useful in analysis of numerous grinding circuit flowsheets of varying-degrees of complexity our technique requires only a consistent process flowsheet with appropriate associated data and arbitrary numbering of streams and equipments. Process topology and user data are decoded to yield a set of "search" or base variables and a set of "calculated" variables in solids, water, particle size and splitter parameter space. This procedure greatly speeds subsequent actual mass balance adjustment calculations since perturing certain variables (e.g., a solids value) does not require recalculation of other variables (e.g., the water balance). Finally, internal search and calculated variables are back converted to User units and used in a Sum-of-Squares criterion. Global minimization of this performance functional then yields a consistent mass balance in a least, squares sense. Weighting factors of choice can be used. Examples illustrate application of these methods.

Jan 1, 1979

By Irwin R. Higgins

CONTENTS A. Introduction B. System Description C. Uranium Pilot Plant Programs D. Summary INTRODUCTION The Chem-Seps Continuous Countercurrent Ion Exchange (CCIX) contactor was first conceived of in 1951 at the Oak Ridge National Laboratory. In the 27 years that have passed, there has been a steady stream of pilot plant process development studies on a variety of processes far too numerous to mention. These have been in the area of water and waste treatment, chemical processing, ore leach treatment, and many others. Many successful production systems have been installed and operated in a variety of applications. A few of these are described herein together with a description of the system and its operational characteristics. SYSTEM DESCRIPTION Chem-Seps Continuous Countercurrent Ion Exchange* is a single loop in which various sections are separated by valves to isolate the sections performing different functions simultaneously. (Refer to Figure 1). To explain the working of a typical unit, let us use a simple configuration used in water softening or demineralization. The resin is moved in small increments by the hydraulic ram action. All solution flows are countercurrent to the flow of resin. The contaminants are removed in the process cycle and the resin is moved in the pulse cycle. A controller in the effluent stream from the treatment section senses the proper limit permissible in the effluent. In the process cycle, the influent enters through a set of distributors, moves downward and leaves through a set of collectors. Simultaneously the regenerant enters through valve RE and leaves through valve WA. The rinse enters through valve RI and its duration is controlled by an interface controller. In the process cycle, valves B, C, and D are closed, isolating all the sections. The duration of the process cycle is 4-10 minutes depending on the contaminants to be removed and is controlled either by a timer if the contaminants are fixed, or a controller in the effluent stream, if the contaminants fluctuate at short intervals. Valve A is open and resin is dropped to the pulse section simultaneously with the removal of contaminants in the treatment section, regeneration of loaded resin in the strip section, and rinsing of regenerated resin in the rinse section. As the resin drops through the valve A, it is backwashed to remove the suspended solids or resin fines. Removal of these fine particles is important to avoid the unnecessary pressure drop in the bed and plugging of the distributors. At the end of process cycle, all flows going into the unit and out of unit are automatically shut off. The valve A closes and valves B, C, and D open in proper sequence. A slug of water enters through valve PU, a predetermined incremental volume of resin moves into the regeneration section, an equal volume of regenerated and rinsed resin enters the bottom of the treatment section, and an equal volume of the loaded resin leaves the top of the treatment section and goes into the resin reservoir section. (The resin movement is plug flow with the Reynold's number being in the laminar region of the curve, and no fluidization occurs during the pulse cycle.) The total time required for proper sequencing of valves and the movement of resin is 20-30 seconds. At the end of this period, the valves B, C, and D close and valve A opens. The solution valves open and normal steady state operation begins. All valves are interlocked and properly sequenced. The resin movement is precisely controlled by sonic devices. Special Features of Chem-Seps CIIX "Downflow"*: The raw feed in the treatment section moves downward and the resin moves upward in the treatment section. The "downflow" of feed overcomes the handicap encountered in "upflow" systems. The "upflow" of feed requires a certain minimum flux during process cycle to lift the bed and keep it

Jan 1, 1979

By Paul D. Tomlingson

An effective maintenance program ensures the adequate availability of production equipment. In turn, this availability enables production to maximize utilization and successfully meet production targets. An overall production strategy is necessary to draw all departments toward this common effort. Only then will production targets be met. From the maintenance view this means that production must comply with PM and repair schedules. Purchasing must supply parts and materials on time. Mine engineering must not overwhelm maintenance with non-vital construction. Accounting must provide correct, timely cost information - - etc. This strategy - - its content, effectiveness and enforcement is the responsibility of the mine manager . . . his alone. This paper describes a successful approach to maintenance within this strategy.

Jan 1, 1979

By C. G. Williams

This paper describes the planning, development and operation of the Highvale Mine to supply coal to both the existing Sundance and the approved Keephills thermal power plants of Calgary Power Ltd. Reference is also made to the company 's Whitewood Mine, which feeds the Wabamun thermal plant .

Jan 1, 1979