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By Alfred Petrick, Roger Dewey

INTRODUCTION Estimation of capital and operating costs for mines and mills has long been a necessary ingredient of feasibility studies, and though many practitioners have successfully used "rules of thumb" and factor estimating methods for many years, a great deal of work is usually required to prepare an initial estimate and/or to update or revise such an estimate. Development of comprehensive mining and milling cost models during the 1970's by the U.S. Bureau of Mines and by Canadian workers (O'Hara and Mullar) provided the basis for feasibility-level cost estimates, but the numerous equations and the apparent rigor and complexity involved in preparing a project estimate using these models has discouraged many, if not most, potential users. Their acceptance was also hindered by a broad misunderstanding by many engineers in the minerals industries who had little exposure to such procedures, and by a possibly inadequate level of explanation or documentation of the parametric approach to cost estimating. The preliminary nature of most databases, geologic/engineering/ and market, available to the engineer or analyst during the early stages of a mineral project is a universal problem. Basing corporate mega-investment decisions or even individual equipment selection decisions on the usual "SWAG’s has always been a risky practice, but in the current business environment, the "guesstimates" are almost 'verboten', and are certainly becoming antiquated with the advent of both the described models, the ubiquitous microcomputer, and the nearly-ubiquitous hybrid species known as the "computerized engineer". Engineers have almost always sought techniques of cost estimation that can be applied and justified when geologic and engineering information for a mining project is preliminary and incomplete. The answer, in some cases, is the "parametric" approach. Equations, based on actual, past experience are developed to relate capital and operating costs to major physical variables such as the rate of mining or milling, with modifiers for such details as mining method, process characteristics, methods of construction, or types of equipment . Cost Engineering Standards Five major types of fixed capital cost estimates have been defined by the American Association of Cost Engineers : 1. Order-of-Magnitude estimate based on previous cost data and minimal knowledge. usually a "ratio estimate" with no better than +/- 30% accuracy. 2. Factored or ratio estimate based on major equipment costs with probable accuracy in the 25 to 30% range. 3. Budget authorization or Preliminary estimate, based on sufficient project data to

Jan 1, 1987

By Arthur E. Lord, W. Martin McCabe, Robert M. Koerner

Following the driving of each new section of tunnel, a prompt decision is required regarding the type and extent of primary support to control convergence and prevent ground failure. Sometime thereafter, the effectiveness of the in-place support must be evaluated over the entire length of excavation. It is therefore extremely desirable to have available a lightweight, portable, easy to install and interpret, cost effective monitoring system to assist with these ground control functions. Such a system can be assembled based on the acoustic emission (or microseismic) technique. This monitoring technique uses a piezoelectric transducer to detect subaudible noises associated with deformation of and/or crack propagation in the soil or rock being monitored. After detection, the signals are then amplified, filtered and electronically counted to provide a quantified assessment of the stability of the structure. In the simplest sense, the absence of emission, or counts, signifies no deformation and thus stability. Conversely, the occurrance of counts signifies instability to varying degrees. The higher the count rate, the greater the instability and the need for immediate or additional support. Described in this paper is an overview of the acoustic emission (microseismic) method, details of its use in a laboratory study measuring convergence and stand-up time in soil, and the behavior of rock samples under similar stress conditions. Field use of the method is recommended for ground control monitoring in tunnels, and some suggestions regarding monitoring procedure are presented.

Jan 1, 1979

By George M. McArthur

General Statement In Montana, as in most states, the requirements for water quality control are becoming more restrictive. What was once accepted practice is now obsolete. The planning and construction of water quality control facilities has taken on new importance. Past practices and new ideas must be examined to meet the challenge. By this paper we hope to encourage the exchange of ideas for the improvement of water quality control techniques. Both the mining industry and governmental agencies should benefit from such an exchange. The following is a summary of the work done and in progress by The Anaconda Company in Butte to provide an improved water quality related to production processes. We invite your comments. Acknowledgements This paper would not have been possible without the efforts of many individuals and departments in The Anaconda Company, including the Environmental Engineering Department, Mr. Frank J. Laird, Director; Montana Operations, Mr. M. K. Hannifan, General Manager, and their assistants.

Jan 1, 1972

By J. Raatikainen, I. Auranen

"INTRODUCTION Sorting technology has advanced rapidly recently. Of the two possibilities particle sorting has advanced especially in introducing new sensors and the use of multiple sensors. The analyzing and computing lead times have also shortened. This has made it a viable technology option for many applications but not for the most important sorting task at a mine; the run-of mine ore. Bulk sorting, however, has advanced in smaller steps, even if the technology has the capability to tackle the run-of-mine ore and offers a high potential to bring mines substantial economic benefits. The sensors, which are useful are restricted by the difficult operating conditions or by the ore characteristics. One of the important features of run-of-mine sorting technology is the major material handling infrastructure required. The capital expenditure can be substantial requiring good understanding of the sortability curve (grade-recovery curve) behavior of the ore. This paper discusses one industrial case, where substantial amount of data measured from the ore feed belt has been obtained. This data was then analyzed for the grade recovery curves obtainable from that ore with sorting. EXPERIMENTAL The XRF measurement used was the FCA device from IMA Engineering (Fig.1). It is a specially built version of XRF designed to operate with coarse material. The measurement was situated at the main belt conveyor taking ore from the primary crusher to the stockpile. The crushed ore was about 80% -300 mm at a capacity of 1100-1200 tph (Fig.2). Measurement times were varied from 30 seconds to a minute. Measurements were obtained over several months.As the XRF is a surface distribution measuring technique, the measurement was first tested upon accuracy and reproducibility against plant daily wet chemical assays and against on-line Courier feed analysis. As there is a varying delay in the ore stockpile, the delay time was used as a variant in the validation. Different measurement times were tested for accuracy. A one minute measurement time gave a total measurement relative standard error, which was both for Cu and Ni below 3%. The results showed also very robust correlations; for Cu, the correlation coefficient was R > 0,94 for one minute measurements over three months and R > 0,86 for Ni respectively. Fig 3. Shows a measurement period of two hours showing substantial feed grade variations (given as relative value; 1.0 represents the mean value during the measurement month)."

Jan 1, 2018

By C. C. Nolan

As the size of the mills get larger and the liners get heavier, abetter means had to be found to handle liner replacement. McLellan Industries is famous for their innovative ideas andwas called upon by an engineering company to develop a mill liner crane. The first crane was developed for Similkameen about 1969 to serve a 32' Ø X 14' Long SAG Mill. The next was Island Copper in Port Hardy, B.C., the same size as Similkameen in 1970. At that time, we took existing hydraulic cranes and modified them to meet the mill configuration. In 1973, we developed a completely new approach for the Southern Peru copper mine at Tacna, Peru for a Ball Mill, size 16'-6" Ø X 19' Long. Each installation required a design study to meet the change in mill size, trunnion size, elevation of floor to feed trunnion, area at the feed or discharge end of the mill. We have supplied self-propelled units as well as lift in. When there is no room at the feed end and if the mill has a trunnion overflow discharge, we have the mill supplier provide a structural frame to span the sump and move the mill liner handier into position with the use of the overhead crane or a mobile crane.

Jan 1, 1999

By Archie M. Richardson, Carl E. Brechtel, Tom R. Kelly, Frank Feero

Recent work to upgrade the ventilation system for the el Mochito lead/zinc mine in central Honduras is discussed. Network modeling and underground measurements were used to evaluate cost-effective alternatives for achieving satisfactory ventilation in a complex and expanding underground operation. Both interim and long-term solutions were implemented to make mining possible under difficult conditions. INTRODUCTION Mining operations at the el Mochito Mine, located in the central highlands of Honduras, Central America, have been virtually continuous since the opening of the mine in 1948. Initially a high-grade silver mine, the mine has expanded along a westward trend of pipe-like orebodies to a distance of roughly one mile from two centrally located shafts. Production of the relatively large San Juan orebody at the western most extension of the mine (Paddock, 1981) led to the introduction of diesel-powered equipment; however, the ventilation infrastructure was insufficient to meet the needs of mechanized mining. A succession of owners/operators in the 1980s allowed the existing ventilation infrastructure to decay to the point that ventilation in the production areas of the mine was very poor. Environmental conditions in some working areas were not conducive to efficient ore extraction because of high dry bulb temperature, high humidity, and diesel emissions. Upon acquiring the mine in 1990, Breakwater Resources of Tucson, Arizona began an aggressive program to refurbish the mine infrastructure to complete extraction of the San Juan orebody and to allow the extension of the mine another 2500 ft to the west for extraction of the Nacional orebody. The program included increasing the capacity of the main ventilation system. This article presents a case history of the process of upgrading the ventilation system in a mine where extensive old workings cause large air leakage. This process has been one of selecting solutions to difficult technical problems that are compatible with the existing mine infrastructure and economic constraints. The initial ventilation system is described in the background section, along with ventilation projections for the mine expansion. Field characterization of the ventilation system for design verification and fan specification is then discussed. The paper describes a series of interim changes to the system to improve ventilation pending completion of new ventilation boreholes. In addition, the temperature/ heat problems in the mine are described. BACKGROUND Initial Condition of Ventilation System The ventilation system is illustrated in Figure 1, which shows the extent of mining with the main ventilation paths superimposed. Early mining around the two shafts opened up vertical connections (stopes and raises) over the entire 2420 ft (737.6 m) of vertical extent, and mining progressed to the west primarily using compressed air and electric-powered equipment. Since ventilation was not a complex problem in the original mining system, the stopes and interconnecting raises were not sealed. The San Juan orebody was much larger than the silver ore zones mined previously, being primarily zinc and other base metals. Its geometry, size, and grade allowed the use of vertical crater retreat (VCR) stoping with diesel mucking and haulage equipment. Its depth, along with the existence of warm groundwater, resulted in a mine climate problem on the lower levels. To establish a complete ventilation circuit, two vertical boreholes (Bonanza Nos. 1 and 2) had been drilled by previous owners from the surface in the vicinity of the San Juan orebody. The system design called for air to be drawn down the intake shafts, across the lower mine levels to the San Juan workings, then up through the San Juan ramps and ore passes to these two exhaust boreholes (see Figure 1). In practice, however, only the Bonanza No. 1 borehole was drawing air through the desired path. Leakage across the old upper levels from the intake shafts, the Caliche tunnel, and from intervening abandoned stopes and raises supplied most of the air flowing to the base of the Bonanza No. 2 borehole. In effect, there were two ventilation circuits in semi-parallel through the mine, of which only one was delivering appreciable air to the San Juan workings.

Jan 1, 1993

By Hsu-Hui Hsing

In several previous studies from this laboratory (1-4) the autophobicity concept of Zisman and co-workers (5, 6) has been extended to gas phase adsorption processes. Low molecular weight alcohols adsorb on silica or alumina to yield novel isotherms in that monolayer adsorption occurs at low relative pressures indicative of strong bonding, but subsequently multilayer adsorption is quite sparse and diminishes with molecular weight of the alcohol. Such isotherms fit no standard classification scheme, however, in the vacinity of monolayer coverage alcohol isotherms give an excellent fit to the Langmuir or B. E. T. , Type I model (7). Monolayer adsorption is relatively irreversible and it is thus possible to prepare these chemisorbed layers and perform subsequent adsorption studies thereon. All indications are that the adsorptive properties on such films is radically different from that on the underlying alumina (4, 8) and in one study finite contact angles for hydrocarbon fluids were observed. One perennial conceptual difficulty is that there is no observed upper bound to the adsorption isotherm which would be necessary for autophobic adsorption to occur. The cause for this has been studied, and it is currently believed that adsorption is complicated by reversible capillary condensation at contact points between powder particles. The evidence for this is that large particle powders give less "apparent" adsorption at high relative pressures and in one very tedious study adsorption on planar sheets, oxidized aluminum displayed similar reduced uptake at high relative pressures when compared with high specific surface area powders (3).

Jan 1, 1974

By Robert A. Jankowski, J. Drew Potts

By implementing various mining practices, mine operators can reduce the respirable dust exposure levels of the roof bolter and the continuous miner operators. The Bureau of Mines has developed a computer program, entitled AMP (Alternative Mining Practices), that gives a quantitative comparison of these reductions. The alternative mining practices and their corresponding reductions in dust exposure levels that are evaluated by the program include a remote control continuous miner, a half-curtain ventilation system, two auxiliary fan-tubing bypass ventilation systems, a modified cutting sequence, a comparison between single-split and double-split ventilation networks, and a comparison between exhausting brattice and exhausting tubing. Descriptions of each of these mining practices are included in the report. The report also includes a hypothetical case study to show how AMP enables mine operators to select the technique(s) that is best suited to their particular needs. AMP is a compiled basic program that has been designed to run on computers with a DOS version 2.0 or later operating system. It includes high resolution graphics (640 X 200 dot positions) designed to run with video boards that are functionally compatible with the IBM Monochrome/ Printer Adapter. Graphic screens are selected from menus, and can be bypassed if the computer is not equipped to handle high resolution mode. Their primary function is to show proper implementation of the mining practices evaluated by the program. This package requires a minimum of 256K bytes of RAM. (Reference to specific products does not imply endorsement by the Bureau of Mines.)

Jan 1, 1987

By J. Kallioinen

Flotation of low grade Ni ores with serpentinite host rock tend to yield low recoveries of Ni With high MgO contents in the concentrate. The recovery can be enhanced by optimizing the sub-processes in the pulp and in the froth. The new OF-100 flotation cell has been designed to be adaptable to different feed fineness (Coarse, Fine). The losses can according the research conducted at Outokumpu's Hitura plant be substantially decreased by OF-100's new rotor and stator design and froth zone geometry. The froth residence time is used to adapt for feed particle size changes measured by the Outokumpu PSI- 200. Models for the behaviour of different particle Sizes in the sub-processes have been developed. The modified froth zone geometry and the capability of maintaining a froth of desired thickness allows efficient washing of the obtained concentrate producing low levels of MgO.

Jan 1, 1994

By D. T. McMurray

INTRODUCTION Shabanie mine, situated some 180 km east of Bula¬wayo, has been a producer of chrysotile asbestos for more than 50 years. The ore bodies occur in serpentinized dunite, which overlies talc-carbonate schist. A zone of relatively competent rock of varying thickness occurs between the schist and the ore bodies, which are gen¬erally less competent. The hanging wall of the ore bodies is economic, and the hanging-wall serpentine carries a variable subeconomic amount of fiber. It is important to note that, in general, the ore body competence is less than that of the foot and hanging-wall formations. Historical After surface operations ceased, cut-and-fill stoping was used to win ore from underground; this was success¬ful until the increasingly stoped-out area caused insta¬bility in the stope pillars and back. Consequently, dur¬ing the early 1950s, a gradual change to cave-mining methods was made, the ore being won by hand lashing in drawpoints, situated in the basement of the stope blocks, and passed through orepasses under gravity to the haulage level some 13 m below. About this time, interest was focused on the sub¬level caving method in use in Swedish iron ore mines: it was felt that it might be applied economically to the Shabanie ore bodies. Accordingly, in 1958, an experi¬mental stope block was laid out in which sublevel inter¬vals and extraction tunnel spacing were 9 m. The tun¬nels (ring drilling drives) were oriented on strike-in contrast to the Swedish system, in which crosscuts that retreat from hanging to footwall are used. The advantages of the method were quickly appre¬ciated by the operating personnel and, despite the in¬evitable teething troubles pertaining to the introduction of any new mining method, it was not long before sub¬level caving was providing a high proportion of the mill feed. The disadvantages also became apparent at an early stage, however, and, from that time to the present, continuing modifications have been made to mining lay¬outs in an effort to improve ore recovery. GENERAL DESCRIPTION OF METHOD The mine is served by a vertical hoisting shaft, in which two skips, a man cage and a service cage, provide adequate capacity for production requirements. The rock hoist is a Ward Leonard control hoist, in which two electric motors drive a common gearbox. The man winder is driven by an a-c motor. Several auxiliary shafts provide secondary egress and intake and return ventilation. Main haulage levels are above (Fig. I a and b). Blasthole fan patterns are drilled by drifters of 100 mm bore, drilling 41-mm holes; when a sufficient strike length has been drilled, a slot is cut in the upper¬most sublevel and the rings are broken into the slot. Initially, a limited tonnage is drawn, since it is essential to ensure that the hanging wall caves behind the retreat¬ing stope face. Once this has been established, maxi¬mum tonnage can be drawn, as described later in this chapter, under the heading "Draw Control." The broken rock is loaded by 0.14 and 0.20-m3 load¬ers into cocopans (rocker-dumping type of tipping truck), which are hand trammed to orepasses, discharg¬ing on the haulage where 11-t electric trolley locomo¬tives haul 3.95-m3 Granby cars to the main shaft bins. As is evident from Fig. 1, the layout is simple, the block is brought rapidly into production, there is a high degree of selectivity and flexibility, and the result is a low-cost high-productivity mining method. DEVELOPMENT Main haulages are developed at 3.2 x 3.2 m, and once the service winze connections have been completed the development of the sublevels is undertaken. The footwall drives are cut first, to obtain access to the block. These ends are of the standard section, 2.4 x 2.8 m, and from them crosscuts at intervals of 70 m are driven through the ore body to the hanging wall. These crosscuts are used to supplement the geo¬logical information previously obtained from diamond core drilling, and they provide additional and more de¬tailed data on fiber percentages and lengths, structural features, and other relevant criteria which are used to build up the geological assessment of the area and to classify it in terms of the geomechanics rock classification (Laubscher and Taylor, 1977). The crosscuts also allow the necessary orepasses to be sited conveniently so that tramming distances from the loading points are not excessive. Development Drilling Once the skeleton development has been completed, the extraction headings are developed at 2.4 x 2.4 m as shown in Fig. 1. Standard development practice is to use crews of a machine operator and his helper, equipped with air-leg mounted jackhammers, to drill rounds of 1.8 m with integral tungsten carbide tipped drill steel. The round drilled is a normal drag round, as shown in Fig. 2, but considerable attention is paid to the drilling of the perimeter holes to use effectively the

Jan 1, 1982

By A. L. Mular

Steady-state material balances calculated from data measured at various locations around process units are useful for many purposes, such as metallurgical accounting, on-line control, and the development of mathematical models. Often it is possible to calculate material balances by several independent procedures when excess measurement information, i.e., redundant data, is available. Clearly, if the data were collected without measurement errors, a theoretical condition never found in practice, all material balances calculated from redundant data would be in agreement. The real situation is that error exists in practical measurements, so that the results of material balances determined from available optional procedures differ. Consequently, best fit computational procedures to improve the material balances by taking measurement errors into account have been formulated. Such procedures can be used also to judge whether measurements are grossly in error due to either faulty observations or defective data sensing devices. A variety of mathematical techniques are available which serve to adjust raw data. The adjusted data permit mass balancing that does not depend on the computational method. To illustrate the need for an adjustment procedure, Table 1 shows a raw data set acquired around the Brenda Mines Ltd. No. 4 grinding circuit shown schematically in Fig. 1. At least seven optional ways to calculate a circulating load ratio may be employed. These are shown in Table 2 where, depending upon the calculation [ ]

Jan 1, 1979

By CAIJUN CHEN, MING ZHANG, ZHENWEI WU, FENG ZHOU, BOQUAN LI, XILIANG ZHANG, ZHOU PU

Traveling grate is a widely used process for iron ore pellet production, but the thermal state of pellets in the drying section is still unclear. In this study, the temperature distribution of the drying section was investigated using computational fluid dynamics (CFD). The results indicate there is a coupling effect in the drying section of a traveling grate. Gas with higher temperature in down-draught drying (DDD) transfers heat to the up-draught drying (UDD) through the gap of the insulation board, and the temperature gradient of the pellet bed changes before reaching the insulation board. The characteristics of pellet bed expected to appear in DDD are displayed in UDD in advance. To verify the model, a contrast experiment was conducted.

By Rodolfo Lagos, Nicos Pfeiffer, Ben Harding, Clayton V. Deutsch

This paper outlines a cutting-edge probabilistic resource modeling workflow tailored to the Round Mountain epithermal gold deposit. The primary goal is to craft a robust and modern simulation process that enables the quantification and comprehension of anticipated discrepancies during mining reconciliation, offering a more precise gauge of expected gold quantities compared to conventional estimation models.

Dec 1, 2023

By YUSONG MIAO, DI WU, XIANRONG YAN, YIPING ZHANG, JIE LIN, KEBIN LI

Rock fragmentation has been a concern in hard-rock mining for the mining industry. Rock fragmentation distribution is an important index to measure the blasting effect. It not only affects the cost of drilling and blasting, but also affects the efficiency of subsequent operations, such as loading, transportation and secondary crushing.

By LUIS ALBERTO TORRES-CRUZ

The Plewes method is relatively well known among geotechnical engineers interested in assessing the liquefaction potential of soil deposits. Furthermore, it was one of the four methods considered in the post-failure investigation report of the Feijao Dam I, also known as the Brumadinho Dam, in Brazil. The inclusion of the method in that report is likely to increase its visibility among geotechnical practitioners tasked with assessing the liquefaction potential of tailings. While the authors of the method duly presented it as a screening method, more recent technical literature contains conflicting views on its reliability. Considering the popularity of the method and that the safety record of tailings dams is a cause of great concern, this study assessed the uncertainty of results from the Plewes method. The study concludes that, even under optimistic assumptions, the results of the method are too uncertain for detailed characterization of tailings deposits.

By D. John Curtis, Gordon S. Crighton, John C. Hester

The six main drives needed to build 55 miles of 4.8 and 7.7 metre i.d. tunnels in 33 months called for careful co-ordination of the design of tunnel boring machines and linings. The implications of these requirements are discussed, with particular reference to the selection of tunnel liners. The solutions adopted are detailed and details are given of the specifications and procedures adopted. The design methods employed for liners are outlined. The construction methods are described and current performance figures given.

Jan 1, 1991

By Thomas E. Howard

Engineers today face a tremendous challenge. Growing population and rising standards of living all over the world are creating unprecedented and accelerating demands on and for all kinds of technology. At the same time, the public is becoming more and more conscious of--and concerned with--the impact of technology on human welfare. The growing controversy over the Supersonic Transport is a case in point, It is clear that the modern engineer must expect to spend a growing share of his time considering this dilemma and evaluating the tools that are available to him, not only with respect to their adequacy for the job at hand but also their capability for meeting the increasingly demanding--and, in many cases, conflicting--requirements of the future. This meeting seems a particularly appropriate place to speculate about the future of mining technology. The summary of the state-of-the-art in surface mining contained in the papers scheduled for the International Conference, together with the recently published Surface Mining volume, constitute not only a major contribution to the technical literature but also form an impressive record of technologic progress. That record, and the advanced technology displayed in the exhibits, and to be discussed in the SME technical sessions demonstrate engineering achievement in which the mining community can justly take pride.

Jan 1, 1968

By A. P. Massmann

Design for economical operation, minimum operating labor and minimum maintenance for labor and supplies are virtually inseparable. Each consideration has a vital bearing on the other. Let us assume that we have the design complete for the most economical operation with a minimum of operating personnel. This means, in addition, that we have the fewest number of operating units consistent with the product requirements for a team or metallurgical coal which we are called on to produce. The number of units will vary greatly, being influenced by design, washability characteristics, tonnage and degree of quality control necessary by reason of producing a product for metallurgical, steam or domestic use. The units involved most ordinarily for any type coal preparation plant will number as a minimum the following: 1. Hoppering R.O.M. Coal 2. Feeding

Jan 1, 1959

By JIACHENG XIE, Juanli Li, Wenyong Du, Tao Cui, Mengyao Dong, Pengfei Huo

The study proposes a dynamic path planning method for mining equipment based on spatial kinematics and real-time correction of coal-rock state identification. The method integrates real geological data and equipment operation information to improve the autonomous cutting of fully mechanized mining equipment.

May 1, 2023

By Elizabeth Anne Semon

The marketability of a large volume bulk material is significantly effected by transportation costs. Hence, it is essential for anyone that deals with minerals or mineral wastes to be able to reliably estimate bulk transport costs. The three types of bulk overland transportation modes that will be discussed are truck, railroad, and pipeline. Contract and private carriage are the two types of truck transport generally suitable for movement of bulk commodities. Contract carriers are those transportation firms that service a limited number of shippers and geographic points, and carry a restricted number of commodities. Private carriage consists of control over drivers and equipment by the shipper, ownership or lease of the transport equipment, and shipment of commodities related to the primary business of the shipper.1 A combination of private and contract carriage may also be used. Depending on the size and regularity of the shipment, rail offers single car, multiple car, or unit train service to move bulk minerals. Both long and short distance pipeline transportation has been shown economically and technically feasible for many different solids such as, mine and quarry products, coal, chemicals, copper concentrates, phosphate rock, and sewage sludge. 2

Jan 1, 1973