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By Terry I. Mudder

Metal complexed cyanides in wastewaters form as a result of interactions of free cyanide with metals present in the wastewater and exhibit varying degrees of stability, toxicity, and treatability. Thiocyanate, a pollutant commonly found in cyanide bearing wastewaters, is formed through the interaction of free cyanide with sulfur containing species (i. e. pyrrhotite) both present in the wastewater. In certain industrial processes, such as the beneficiation of gold and silver, cyanide is an essential reagent. Since free cyanide, complexed cyanides, and thiocyanate are potentially toxic to humans and aquatic organisms, these compounds must be removed from wastewaters prior to their discharge into surface or ground waters serving as potential potable water sources, marine or fresh water habitats

Jan 1, 1984

By S. Makni, F. Wang, A. Rosa, B. Klein, C. Saud, R. McIvor, A. Kumar, C. Gagnon, D. Gong

In response to the Natural Resources Canada “Crush It Challenge”, Corem partnered with the University of British Columbia and led a project to develop and demonstrate the use of high-pressure grinding to replace the workhorse of the industry, wet ball milling, with high-pressure grinding. Two plant case studies, a small tonnage gold recovery operation, and a large tonnage copper recovery operation, were carried out. Plant circuit performances were compared to pilot plants operated on the same circuit feed set up to represent the novel equipment and flowsheet that was developed to produce separation circuit feed. Numerous obstacles to industrial implementation of highpressure grinding in this role which were recognized at the outset of the work were successfully addressed. In both cases, comminution equipment energy savings alone, and final stage total circuit energy savings, exceeded 60% and 50%, respectively. Downstream mineral separation performance was the same or better.

Mar 2, 2022

By Kadri Dagdelen, Lee J. Zamalloa

Modern surface mining operations are always seeking for minimizing costs while increasing production rates, among the optimization of other specific tasks that have an impact on the economics of the mine. Specifically, the cost related to the production fleet operation (shovels and trucks), the maximization of the overall performance of the equipment, and the uncertainty to meet the goals at the end of the operational time horizon are always topics of discussion with no apparent simple answer. On the other hand, truck-shovel dispatch systems based on Linear/Non-Linear Programming Algorithms have been in operation since mid-1990’s with different levels of success depending on the correct arrangement of operational resources and the complexity of the mine, but having in most cases some important shortcomings related to the way these systems deal with the decision making of the operational objectives, being necessary to have a human agent (dispatcher) to guide the system when the operational parameters change, creating the possibility of suboptimal decisions due to the human factor. This paper discusses the implementation of a novel methodology based on AI techniques to assist in the decision making of the best goal selection when operating a Fleet Management System, taking into consideration time periods as variables and integrating them as part of the analysis, providing a better overview of the forecast plan completion with respect to previous production numbers and evaluating best alternative to handle future events to minimize losses or maximize profits, resulting in better truck assignments, minimum shovel idle times and overall productivity improvements of the operation at the end of the time horizon.

Jun 25, 2023

By Hans W. Schreiber, Rudi P. Fronk

IINTRODUCTION Widespread activities in precious metal projects in the western U.S., Australia, Canada and other parts of the world have opened a flood-gate of consul- ting expertise to aid and accelerate the development process. Utilization of outside expertise has taken on many shapes and forms that have in some instances actually "decelerated" or "retarded" project progress. Small mines are especially vulnerable to apparent false-starts and misdirections, be- cause of their generally lower capital commitments, shorter lead times and limited mineral in-house expertise. Mine development for small mines requires fundamental approaches by knowledgeable consultants that are timely and cost-efficient and in line with accelerated development, if they are to be successful. The development process is obviously that process which brings a given deposit into production. Or, put another way, the process comprises all activities, decisions and commitments which take place from the moment an authorizing entity (usually a board of directors) com- mits to development and production (usually on the basis of a final feasibility study) until the first ton of ore is processed to a saleable product. The activities, decisions and commitments over a development period--such a period can range from as short as a year or less to as long as 5 years or more--are many, highly varied, and often interdependent. The development period is also particularly risky, not only because of the varied and interdependent decisions and commitments, but also because of capital demand amounts. Delays in the development process are thus particularly costly - sometimes so costly that small and often undercapitalized mining companies, failing to reach production and a cash flow within a planned period, face, if not bankruptcy, at least severe equity dilution or particularly onerous and restrictive covenants on additional loan financing. It therefor behooves especially small mining companies to examine ways to accelerate development. By doing so, the development process may actually meet planned scheduling. The development process generally comprises environmental impact studies completion, final engineering, permit- ting, financing, equipment and machinery procurement, contractor selection and construction -- such activities largely being overlapping. The paper addresses what the small mines operator or the small mining company can do with consulting firms to accelerate development. By "small mines" is meant open pit operations of 2000 tpd or less and underground operations of 500 tpd or less. By "consulting firms" is meant entities who do not possess an engineering and construction capability, such as Derry Michener Booth & Wahl, Pincock Allen & Holt, Behre Dolbear-Riverside, etc. The paper concentrates on gold projects. PRIORITY CONSIDERATIONS Of the items listed in the previous paragraph which comprise the development process, the two which most consistently appear to cause delays are permitting and financing. The former's delays can be due to political overtones and thus the delays may not be reasonably foreseeable. The latter's delays are most often due to inadequate preparation, and thus are avoidable. For the reason stated above, and considering that financing is fundamental to all of the activities which constitute the development process, the paper focuses on financing as the priority item which can most assist in accelerating or retarding the development process.

Jan 1, 1987

By Paul M. Jr. Musgrove, Thomas D. Henderson, Mark B. Nesbitt

Introduction History. Vat leaching operations began in October 1953. Since that time, about 4.2 million tpy of oxide ore averaging 0.72% copper have been treated. Economic conditions, improved operating methods, and increased production facilities have allowed the ore grade to be reduced over the years. The average ore grade was 0.84% Cu during the 1950s, 0.68% in the 1960s, and was 0.50%, or below, during the 1970s. By 1961 some 35 million tons of oxide ore had been removed from the mine, exposing the underlying sulfide ore body. A 5,000¬ tpd concentrator was constructed and put into operation to treat the sulfide ore in November 1961.1 The original crushing plants were used for both oxide and sulfide ores, but by adding a third crushing shift, six days per week, the total ore treated was raised from 12,000¬17,000 tpd. Dwindling oxide ore reserves prompted expansion of the concen¬trator to 13,500 tpd in 19672 so that exhaustion of known reserves of both ores would occur at nearly the same time. Again, no additional crushing facilities were added during the expansion. It was anticipated that the treatment rate of oxide ore would be reduced from 12,000 to about 9,000 tpd. Instead, the addition of two crushing shifts per week and improved operating procedures have increased the treatment rate to 26,500 tpd-13,000 tpd to the leaching plant and 13,500 tpd of sulfide ore treated in the concentrator. Geology. The Yerington ore body is a disseminated porphyry copper deposit where a quartz monzonite porphyry intrusive contains the major part of the ore values, with some copper mineralization found in the adjacent granodiorite and other rocks. The geology of the Yerington Mine was aptly described by J. R. Wilson (Mining Congress Journal, June 1963). The ore body is one of few porphyries where geologic conditions were optimum for the formation of an important concentration of oxidized ore with minimum economic effect of secondary enrichment. The principal mineral in the oxidized zone is chrysocolla, which irreg¬ularly occurs dispersed throughout the rock and as narrow seams along fissures. The grade of the ore in the oxidized zone is generally the same as in the primary sulfide below, although some massive concentrations do occur as vein-like fissures and as filling of breccia zones. Cuprite, tenorite, melaconite, malachite, and azurite occur but are not abundant. Lying between the chrysocolla horizon and the primary sulfide is a transition zone whose limits are irregular but seldom with vertical range exceeding 50 ft. Chalcocite, cuprite, melaconite, native copper, and chrysocolla occur superimposed on the primary sulfides. Chalcocite replacement of chalcopyrite, as well as secondary chalcocite, do occur. The primary sulfide minerals are chalcopyrite and pyrite, with minor amounts of bornite and covellite. There are no appreciable gold or silver values, and molybdenite has rarely been noted.

Jan 1, 1985

By P. Roghanchi

The difficulties associated with performing direct tensile strength test lead to development of indirect test methods, which are simple, less expensive, and less time consuming. Splitting tensile strength test (Brazilian test) is performed by applying two opposite concentrated compression load across the diameter of a disk specimen until the failure occurs along the loaded diameter. Theoretically, the load is applied over a small width as a line load, however, by increasing the load, the contact width between the loading jaw and the specimen increases as the specimen deforms. The main aim of this study is to investigate the stress distribution in (1) regular Brazilian specimen in traditional tests with loading jaws and (2) flattened Brazilian specimen with flat platens. A series of the splitting tensile strength tests and uniaxial compression strength tests are performed on three rock types: rhyolite, basalt, and limestone. Further, Numerical modeling studies are performed to investigate the effect of the contact between jaw and Brazilian specimen on stress distribution in the disc specimen based on the results from the experimentation. Finally, the stress distribution in the traditional Brazilian test specimen and the flattened end Brazilian test specimen were compared.

Feb 23, 2014

By James Charles McHardy

During 1980-2000, the Florida Phosphate industry will deplete its Central Florida Polk County reserves. As a result, new mines will be built in the ?so-called? Southern Florida Reserve areas. These new mines will require new mining, beneficiation and tailings disposal technologies in that the southern reserves have different characteristics than have been encountered in the Central Florida deposits furthermore, government regulations will necessitate the development of new tailings and clay disposal methods. This paper reviews these challenges for the phosphate industry and how these opportunities can be expected to effect plant design and plant operating practices.

Jan 1, 1981

By Liming Yuan, Wei Tang, RICHARD THOMAS, DAVOOD BAHRAMI

The location and orientation of fire suppression nozzles is critical in suppressing mobile mine equipment fires. To effectively suppress such fires, optimization needs to be considered for the suppression nozzle locations and discharge angle. Detailed experiments were conducted to study the effectiveness of fire suppression agents with different nozzle configurations. A variety of fire suppression systems including dry chemical, wet chemical, dual agent (dry and wet chemical), carbon dioxide, and water mist were used. Six types of nozzle locations were studied, and their fire suppression effectiveness were compared. Results from this work highlight potential considerations when determining effective fire suppression designs.

Feb 1, 2025

By D. E. Healy

In 1989 an Autogenous Grinding Task Force was formed to study potential autogenous grinding circuits to replace the existing semi-autogenous grinding at the National Steel Pellet Plant located in Keewatin, Minnesota. Pilot plant test work performed by the Hanna Research Center (HRC) indicated that fully autogenous grinding with pebble extraction and crushing could feasibly be incorporated into the existing circuits. In 1990, one of the existing ten grinding mills was converted for an in-plant test incorporating pebble crushing. The full scale test work conducted in 1990 and 1991 verified the pilot plant projections with primary mill throughputs from the fully autogenous circuit utilizing pebble crushing being greater than throughput of the existing SAG mill circuits.

Jan 1, 1992

By Robin G. Burn

An orebody is a mining company's greatest asset; without it there can be no mine. Yet, by the very nature of the industry, orebodies are wasting assets whose progressive depletion must be balanced by continued discovery of new ore to enable both the industry to survive and the raw material demands of society to be satisfied. Although mineral exploration may be viewed as any activity which leads to the creation of investment opportunities in the mineral industry, ultimately new ore can only be generated through discovery. The process leading to discovery may involve a long chain of independent and interdependent events. It will almost always be repetitious and require the examination of many prospects or trials before the goal, a commercially successful discovery, is satisfied. Long term survival of a mining company through regular discovery of new ore, requires a continued commitment of funds and effort to mineral exploration.

Jan 1, 1984

By Peter Zhang, I. BERK TULU, Morgan M. Sears, Heather Dougherty, Ted Klemetti, DYKE MARK VAN, Gabriel S. Esterhuizen

Ground falls in longwall gateroad entries remain a concern in modern longwall operations. The gateroads are subject to changing horizontal and vertical ground stress induced by longwall extraction. These stress changes can result in failure of the strata around an entry leading to large deformations of the entry roof, floor and ribs. The gateroad support systems are required to control the failed strata while maintaining safe access to the longwall face and unimpeded ventilation. This paper presents research conducted to better understand the stability issues in gateroad excavations and to develop procedures for evaluating support and layout alternatives for longwall gateroads. Using the results of a field-monitoring program and numerical model analysis of case histories, a conceptual model of gateroad support needs was developed. The conceptual model formed the basis for developing a set of equations that can be used to estimate likely roof sag and support loading for given roof geology and longwall-induced loading conditions. The developed equations were used to compare predicted gateroad stability to field study results, showing satisfactory agreement. The calculation procedures are used to demonstrate their application in assessing support alternatives at a case study mine. It is concluded that the developed analysis procedures provide realistic assessments of likely ground stability and can be used to evaluate alternative gateroad support systems at operating longwall mines.

By Joseph W. Murray

GENERAL DESCRIPTION The undercut-and-fill method of mining was devel¬oped by Inco Ltd. in the Sudbury district of Ontario, Canada, to deal with abnormal ground conditions en¬countered in pillar recovery. Although it is now some¬times utilized as a primary stoping method, the main application continues to be in pillar recovery. In any mining method in which fill is employed, pillar systems tend to fail gradually as stoping pro¬gresses. As a result, most pillars are completely destressed and broken by the time pillar recovery com¬mences. Control of the back in such ground becomes a major problem and is contingent on local conditions as they exist in the workplace. In most mining methods, operations such as drilling, blasting, slushing, and tim¬bering can be readily assessed and the mining cycle pre¬dicted. However, in the mining of pillars with overhand methods, the time and effort required to maintain a safe back varies to a quite unpredictable extent. For years the square-set method was the most effective means available for pillar recovery. However, in areas where bad ground was encountered, square setting was a costly and sometimes hazardous method. As the tonnage mined from pillars increased, it was apparent that a safer and less expensive approach was required. Past experience indicated that the best way to deal with broken ground was to mine from the top down. Top slicing systems were examined but were rejected due to the major sub¬sidence which resulted. However, ideas generated from studying top slicing finally developed into the present method of undercut-and-fill. This method, which has almost completely replaced square-set mining, exhibits reduced costs, has improved safety and efficiency, elimi¬nates the factor of bad ground overhead, and allows a closer prediction of the actual mining cycle. The result is more accurate scheduling of pillar production. This factor is of increasing importance as the tonnage from pillar recovery increases in relation to total mine pro¬duction. Undercut-and-fill is a method of extracting a block of ore by mining successive cuts, working from the top down. After a cut of ore is completely mined out, lami¬nated timber stringers are constructed along the sides for the full length of the cut. Round logs covered by a wire screen are laid across the stringers to form a mat. The cut is then tightly filled with hydraulically placed cemented sand fill. Mining is then resumed on the cut immediately below the mat. As the cut advances, the timber stringers are supported by round timber posts seated on the solid bottom of the cut. Drilling, blasting, muck removal, and timbering operations are repeated until all the ore in the cut has been removed. At this time, another log mat is laid and the cut is filled. This sequence is continued until the block has been mined from level to level. The basic principles of undercut-and-fill can be adapted to a variety of mining situations. It can be applied transversely or longitudinally in ore bodies of varying width, that dip from vertical to flat lying. The method can be applied using conventional equipment or it can be mechanized through the introduction of small diesel or air-powered units for muck removal. Undercut¬and-fill is a selective method, ore recovery is high, and dilution can be controlled to an acceptable level. How¬ever, it is a high-cost, labor-intensive method and as such its use is restricted to areas where the less costly meth¬ods of ground support are not successful or where bulk mining methods are not adaptable. One basic require¬ment for undercut-and-fill is an adequate source of hydraulically placed cemented sand fill and a dependable sand plant and distribution system. The success of the undercut method is to a large degree dependent on the competence of the fill placed in each of the successive cuts. In theory, the weight of the fill on each cut is supported in part by the cemented sand itself as it binds to the rough sides of the mined¬out areas next to the pillar walls, and in part by the tim¬ber support. As mining progresses beneath the mat, a downward pressure is exerted within the cemented fill and a voussoir arch develops above the mat, across the full width of the cut. This supports the major portion of the load. The log mat and posts have a double func¬tion in that they carry the remaining weight not sup¬ported by the voussoir arch and in addition prevent any pronounced settling of the fill mass. Excessive settling breaks the bond between the cemented fill and the pillar walls, eliminating the arch effect and transferring the entire load to the mat and posts. The theory of a voussoir arch forming within the cemented sand fill mass is substantiated in practice by the fact that pillars have been mined from level to level with no more significant weight on the bottom cuts than was experienced on the top cuts. In effect, as mining progresses downward, there is no additional transfer of weight from previous cuts to the cut being mined. APPLICATION OF UNDERCUT-AND-FILL Because of their size, the Sudbury ore bodies are, for the most part, mined transversely and, to a considerable extent, by filling methods. Transverse stoping blocks are designed with 9-m (30-ft) crown pillars established be¬low each level and with stopes separated by vertical rib pillars. In past years, floor pillars were left above the level as well, but present practice favors the silling of primary stopes at the base of the rail. Rib pillars are designed to meet the ground control requirements of the primary stoping methods. Local conditions and past experience generally dictate their size. Historically, since the square-set method was to be used for recovery, rib pillars were designed to accommo¬date standard 1.7-m (5'h-ft) sets. A typical rib pillar is four sets wide [6.7 m (22 ft)] when located between two square-set or cut-and-fill stopes having a width of 8.4 m (27.5 ft). When the undercut-and-fill method was de¬veloped, a mat width of two sets was adopted. This di¬vided each cut of the common four-set pillar into two

Jan 1, 1982

By Christopher Mark

Area wide pillar failures, or “squeezes,” were once fairly common in underground coal mines. While rare today, they do occur. This paper describes seven case histories that were investigated by MSHA Technical Support. They occurred across the range of US coalfields, in both longwall and room-and-pillar mines, and involved development mining, pillar recovery, longwall panel extraction, and multiple seam interactions. Significant lessons from each are discussed.

Jun 25, 2024

By T. W. Lloyd

Accurately predicting coal seam discontinuities is an important part of modern underground coal mining. Conventional methods of coal exploration sometimes fail to identify rapid variations that may seriously impact mining and quality. Radio Imaging Method (RIM), a geophysical approach to mapping longwall panels, has been developed where electromagnetic signals are transmitted and received though the coal block. The amount of signal loss for a given distance, when compared to the coal seam characteristics, identifies anomalous areas. A comparison of RIM predictions with actual geologic conditions encountered during longwall mining demonstrates the accuracy and usefulness of the data.

Jan 1, 1986

By Biswajit Samanta, Pranjal Pathak

This article presents a customized genetic algorithms (GA) with new crossover and mutation operators to solve the short-term production scheduling problem of multi-mine mineral value chains (MVC). The preceding problem consists of determining the extraction sequence and destination allocation of blocks from all the mines collaboratively while closely meeting the quality and quantity requirements of the processing units subject to relevant technical and operational constraints. The short-term production scheduling is carried out at shorter scales wherein the operations are modeled in a great detail with large number of constraints. This makes the industry-scale instances of the problem computationally intractable for standard mixed-integer programming (MIP) solvers. Thus, a GA-based heuristic approach is developed to obtain nearoptimal solutions to the large-scale instances of the problem in a reasonable amount of computational time. Computational experiments show that the developed GA-based method is a promising way to handle industry-scale instances of the problem. Moreover, the sensitivity analysis on various parameter combinations of crossover and mutation operators indicates that the customized global mutation operator, when used in combination with the customized crossover operators, took on average 12.5% less time than the customized local mutation operator to converge to a solution.

May 21, 2022

By K. C. Ko

The pit slope angle dictates the economics as much as the depth of the pit. Steeper slope provides more ore reserves, at the same time, the steeper slope presents higher probability of instability and associated maintenance problems. While large open pit mines mostly belonged to nonferrous metals in the past, other mines are becoming deeper as the mine economics permits deeper mine operations, Such minerals include coal, uranium, phosphate, and others. There are some basic differences between an open pit slope and slopes that are encountered in usual civil engineering work such as dam embankments and highway cuts (Figure 1). The stability of the civil engineering slope changes continuously with time due to ground-water movements, weathering, etc. But the stability of the open pit slope (Figure 2) changes more drastically day-to-day or even hour-to-hour, These changes occur because not only is the slope geometry being continuously changed by the mining operation, but also different geology and geologic structures are being exposed on the slope. Even when a particular slope area is not presently being mined, excavation in adjacent areas can affect the general condition of that slope area. A pit will experience some slope failures during its life.

Jan 1, 1977

By E. F. Nelson

The 1970s have witnessed the completion and successful start-up of several new non-ferrous metallurgical operations in the Western Hemisphere which utilize large semi-autogenous mills in the grinding circuits. These facilities' are arranged either as single-stage circuits, or in a two-stage configuration n with ball mills. Another major concentrator, designed with an autogenous mill, ball mill, crusher (ABC circuit), is presently under construction and scheduled for start-up late this year. The large primary mills employed in these circuits have high diameter-to-length ratios and are designed to accept a very coarse feed, i.e., the product of a primary gyratory crusher, or run-of-mine ore. This characteristic obviates the need for a secondary-tertiary crushing plant or, in selected cases, even a primary crusher, which is a basic requirement in feed preparation for conventional grinding mills. This paper discusses three such grinding circuits and, presents operating data for two of them: the Island Copper Mine of Utah Mines, Ltd., near Port Hardy, British-Columbia,- and the Pueblo Viejo Cyanide Leach Plant of Rosario' Dominicana, S.A. on the island of Hispaniola. The design criteria for the Cyprus Bagdad Concentrator under construction' at Bagdad, Arizona, are also included. Data pertinent to these circuits are tabulated in the Summary.

Jan 1, 1977

By K. L. Gabby

Since mercury is chemically similar to silver, it tends to dissolve along with the silver and gold during cyanide leaching. Ultimately mercury is recovered along with the silver as mercury/silver/gold amalgam which requires retorting to remove the mercury prior to refining the silver and gold. A possible approach to separating mercury from silver and gold in cyanide leach solutions is to take advantage of the replacement reaction between [Hg(CN)4]2- and Ag2S, which results in precipitation of mercury as HgS while dissolving silver as [Ag(CN)2]-. Results are presented demonstrating that mercury can be removed from cyanide solutions with high effectiveness, while, at the same time, recovering additional silver in solution. Initial experiments with static flasks in the laboratory showed a high degree of mercury removal could be achieved. Results using actual metallurgical plant products in bottle roll tests indicate the process works on a larger scale, but also results in significant gold losses if the solution being treated has high gold levels, such as carbon stripping pregnant eluent solution.

Feb 23, 2014

By A. W. Khair

This paper presents a strength formula for the design of coal pillars. The formula is derived from results of laboratory experiments, studying the effects of primary factors influencing the strength of the coal specimens/pillars. Five primary factors, significantly affecting the strength characteristics of the coal specimens/pillars were identified and extensive laboratory investigations were carried out. The five primary factors were namely 1) coefficient of friction between loading platens and coal specimens, 2) elastic properties of the loading platens, 3) coal specimen geometry, 4) coal specimen size, and 5) rate of loading. Four to nine levels of variation for each primary factor on the strength of coal were studied. An empirical formula, implementing the variation of the primary factors affecting the strength of coal specimens/pillars, has been developed. A comparative analysis of the results provided by this and some well known formulae is also presented.

Jan 1, 1994

By Y. J. Wang

In the mathematical analysis of electrical networks, the electrical devices are often modeled using idealized network elements such as current sources, voltage sources, and resistors. An analogous approach is outlined here for the modeling of mine ventilation networks, which can be considered nonlinear resistive networks. The modeling technique proposed uses five basic (non-reducible) elements: the H element, the Q element, the D element, the R element, and the N element. These elements are characterized by their (a) pressure gain or loss, (b) air quantity, and (c) pressure-versus-quantity relationship, each of which may be known or unknown. In addition to the definition of the basic elements, the necessary conditions for which there exists a unique solution are discussed. This is performed using theoretical developments derived from electrical circuit theory and illustrated in example networks. The proposed procedures are useful in the solution and characterization of mine ventilation networks.

Jan 1, 1996