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By R. E. J. Putman

Rising demand, artificial market conditions, increased overseas competition, and a fluctuating international situation have brought ever increasing pressures to bear on the nonferrous metals industries, and particularly in the United States, Although capacity is being expanded by the construction of new plants and the development of new ore deposits, increasing attention is being focused on the possibilities of improving the performance of the existing installations and of getting the most out of the new equipment now installed. At first sight the task appears formidable. Based on raw materials which are nonhomogeneous, and possessing characteristics which are hard to quantify, even the processes by which the values in the ores are concentrated and extracted are in many cases not fully understood. The necessary analytical equipment and transducers demanded by these industries have also been slow in development and their earlier absence has hampered the application of modern control technology, This is reflected in plants which have been designed for ease of manual rather than automatic control, and in which the application of control techniques at first sight appears economically prohibitive. However, in any industry progress often comes only in small steps, It may be that by the wise choice of limited objectives and perhaps only partial solutions, substantial improvements in plant performance can be implemented now, and with only a modest investment. If modern control equipment is chosen of the type which can be reprogrammed, advantage can be taken later of more sophisticated solutions as the technology of the industry advances. This task of improving plant performance can and must be commenced even though our knowledge is incomplete. The experience of a related mineral consuming industry, with similar problems in the past, can be encouraging.

Jan 1, 1967

By G. Daws, N. Woodward, A. Oxley

Many mines around the world have been seeking ways to achieve sustainable developments and reduce operating costs. Rockbolting is a technology that needs to be considered as it provides a means of reducing operating costs and improving safety whilst allowing improved production to be achieved. The problem is how to introduce this technology in the knowledge that a successful outcome will be achieved. Rockbolting is a complete package and not only includes the consumable materials but also training, production of Codes of Practice, Monitoring Schemes, Refinement of Design and long term management of the systems. This paper explains the methods that have been applied for its successful introduction and discusses several case studies. It does not give fine detail but is intended to allow decision makers to gain a basic understanding of the method used to successfully introduce rockbolting to a mine.

Jan 1, 2011

By Kristian E. Waters, Hao Ma, Haibei Wang

Over the past few decades, ELM based separation processes have become an attractive and efficient way to remove toxic metal ions from the various aqueous waste effluent streams. This paper provides a comprehensive review of the generation, characterization, and applications of emulsion liquid membrane (ELM) from the more than 280 papers and books published so far. First, the ELM formation and the transport mechanism are introduced. Then the ELM stability and emulsion splitting methods are discussed in detail, followed by the latest research and findings of the extraction of copper and nickel using the ELM technique during the past few decades, including a thorough discussion of ELM operating parameters that impact ELM stability and extraction. In addition, the kinetics study and industrial application of this technique for the removal of coppers/nickel solutes are also presented and discussed.

Jan 6, 2023

By Arif Emre Dursun

The aim of this study is to present hazard classifications of location, type, and causes of accidents in Turkish coal-extracting industries using Analytic Hierarchy Process (AHP). In this study, fatal mine accidents were analyzed by pairwise comparison and weighting, and the most important hazards were identified by determining the effective main and sub-criteria among location, type, and causes of accidents. For this purpose, 1st, 2nd, and 3rd level hazards consisting of location, type, and causes of accident were determined by using the data of fatal mine accidents in underground and open-pit coal mining in Turkey between the years of 2010 and 2020. And then the data were categorized, and each category was prioritized using the AHP technique. According to the weights calculated by AHP, the most dangerous main criteria at 1st level hazards in underground coal mining is ventilation operations (0.363), while the most dangerous sub-criterion at 2nd level hazards is firedamp explosions and other gas accidents (0.409). On the other hand, in open-pit coal mining, the most dangerous main criteria at 1st level hazards are slopes and benches (0.575), while the most dangerous sub-criterion at 2nd level hazards are landslides (0.429).

Dec 30, 2023

Satisfactory means of managing growth in boom towns is a concern in mineral, energy and other large industrial developments. This session will include examinations of socioeconomic impacts, major issues such as social disruption, case studies, methods of managing growth, and concerns involving Indian lands. Innovative methods will be offered to assist the public, government, and industry in finding solutions to managing growth in rapidly growing communities.

Jan 1, 1979

M.S. Azun I have many objections to the content of the author's paper. Before discussing it, however, I would like to repeat the property of semivariogram function. Second order stationary properties of regionalized variables (ReV's) such as semivariogram function y(h) are perfectly known in geostatistics. Also, the kriging equations in the language of mathematical statistics using second order stationary properties are well understood. However, the way to use the sample (estimated) semivariogram function in any one of the kriging procedures is vague. The sample semivariogram function is given as follows: [ ] where N is the total number of samples, Xi is the sample value at the i - th location, Xi+h is the sample value at the i + h - th location, and h is the distance among the samples. An estimation variance of sample semivariogram function of first lag is smaller than that of higher order lag. The theoretical semivariogram function reaches the variance of samples asymptotically. But this is not easily observable because of the larger variation involved in the estimate of semivariogram function. In general, an estimation procedure is done for h = 0, 1, 2, up to the greatest integer less than N/2, even though sample semivariogram function can be computable through N-1. After estimating semivariogram function, the critical question of how to model sample semivariogram function arises. As seen in the above equation, sample semivariogram function is discrete and can be smoothed by the model being selected. Therefore, modeling of sample semivariogram function is the most important step in geostatistics. It not only smoothes a discrete function but also affects the results of the kriging procedure. When the only aim is to model the semivariogram function, which is the basic point of the author's paper, one can employ any fitting techniques, such as curve fitting, or any arbitrary functions, which are called submodels in the paper. The term "arbitrary function" is used rather than "submodel" because there is no basic understanding of developing them. The author suggests that the sum of those submodels can also be used for the modeling of sample semivariogram function. The combination of any arbitrary functions brings many problems instead of giving an insight of the domain structure considered. The author used two arbitrary functions and the nugget effect in response to sample semivariogram function (Fig. 10). For the same example, he stated that the parameters involved in the mixed arbitrary function model can be accepted when the discrepancy between sample semivariogram function and the model is small visually. For verifying the fitting behavior of any selected model, one should not be contented with the visual satisfactory. Some statistical measure such as goodness of fit has to be used. The author's practice is no more than an exercise in curve fitting without any fundamental understanding or conceptualization of the underlying physical mechanism. Furthermore, the selection of any model is not an easy task if the purpose is the search for the "best" response to the observed second order properties of ReV's. I suggest that the Markovian model (Azun, 1983), on the basis of a theoretical understanding of underlying mechanism, which gives more information about the occurrence of regionalized variables, is used to respond all properties of ReV's. There are a lot of problems for modeling of one-dimensional sample semivariogram function. Thus, it is not appropriate to go to higher order dimensional sample semivariogram function modeling. In the meantime, I would recommend that one can connect the values of standardized sample semivariogram function rather than simple values of semivariogram function in the two-dimensional estimation. The standardized values can be computed in dividing the semivariogram function value by the number of sample pairs involved in each lag regardless of the directions. In conclusion, geostatistics is an interdisciplinary area in mining that uses the principles of mathematical statistics. Thus, it should not violate any probabilistic and statistical rules. When Matheron was developing the theory of geostatistical study in the early years of geostatistics, many mining people had a reservation accepting the geostatistical tools. However, this does not mean that we, the geostatisticians, might try to convince those people using some "strange" tools or rules as some authors implied (Baafi and Kim, 1984). Instead, we have to develop and explain the geostatistical tools staying only in the framework of statistical concepts and properties. References Azun, M.S., 1983, "Stochastic Process Modeling of Spatially Distributed Geostatistical Data," Columbia University, Ph.D. Thesis. Baafi, E.Y., and Kim, Y.C., 1984, "Discussion - Comparison of Different Ore Reserve Estimation Methods Using Conditional Simulation," Mining Engineering, Vol. 36, No. 3, p. 280. Reply by J.M. Rendu The interactive method proposed by Rendu allows practitioners to develop semivariogram models that take into account not only the numerical information obtained by sampling, but also highly significant additional information that often cannot be quantified. The geology of the deposit - including hypotheses concerning its genesis, sampling methods, assaying methods, and mathematical methods used to calculate the semivariograms - all have an influence on the numerical results obtained and on how these results should be interpreted. If all the information concerning the spatial distribution of values in a mineral deposit was contained in the sample values, it could be argued that statistical techniques alone would produce optimum models. However, this is rarely, if ever, the case. Methods that allow the user to take into account his experience and his geologic understanding of the deposit should not be rejected for the sake of theoretical statistical purity.

Jan 11, 1985

By John H. Bassarear

Large scale equipment can reduce both capital and operating costs for concentrators. Large semiautogenous mills can be effectively utilized in big mills but they must be kept fully loaded to take advantage of their great potential. With varying hardness ores, controlling throughput is difficult and careful design must take this into consideration. A variable speed drive on large primary semi- autogenous mills can overcome this problem to a considerable degree. This paper provides a base case using standard accepted technology of two wound rotor single speed 4,476 kW (6,000 hp) motors to drive the primary SAG mills of a hypothetical 36,364 mtpd (40,000 stpd) typical porphyry copper mill. One other fixed speed drive will be described and five variable speed drive configurations will be reviewed. Capital costs for the entire grinding circuit, from crusher product to flotation feed, were developed for the base case and each of the alternatives. A technical evaluation comparing the advantages and disadvantages of each drive is presented.

Jan 1, 1985

By Robert D. Deister

The Buick Concentrator represents state-of- the-art technology in automated digital process control operation utilizing a centralized computer. The process control strategies implemented at Buick are no longer new to the industry but the successful application of these developments is still rather unique in North America. Productivity improvements, determined to be a direct result of the process control developments, are the primary topics of this paper. Emphasis is placed on metallurgical benefits including both lead and zinc recoveries, lead concentrate grade and reductions realized in reagent consumptions, operating costs and personnel. The productivity improvement discussion is preceded by a concentrator description and a brief history of the process control development at Buick. Current control strategies and future applications are presented afterward.

Jan 1, 1985

By S. R. Titley

Carbonate-hosted massive sulfide (60%) ores of the western Cordillera compose a distinct class of deposits although they vary widely in detail and perhaps genesis. They have been important sources of lead, zinc, copper, silver, gold, and more rarely, molybdenum, tin, and tungsten in complex mineral assemblages usually dominated by pyrite. They occur as concordant stratiform bodies, as discordant mantos, and as chimneys crossing tens of meters of section; many are associated with faults and breccia bodies. Environments of formation of carbonate hosts range from craton basin to shelf with ages from Proterozoic to Cretaceous; in many instances, dating has shown mineralization to be significantly younger than hosts. Temperatures of formation determined for many deposits are high (200°-400°) and many districts are centers of igneous activity manifested by dikes, sills, and stocks. However, direct cause and effect relationships have been difficult to establish. Many important problems attend the renewed interest and study of these deposits and include questions of metal-sulfur sources, metallization processes and histories, geological controls at all scales, and relationships to broader patterns of metallogeny.

Jan 1, 1985

By T. E. Masoner

The struggle to get the minerals industry seriously interested in high intensity wet magnetic separation* has been difficult. The high capital cost of the equipment, the lack of basic research, the lack of operating information, and the diverse characteristics associated with different minerals and ores have prevented the development and use of HIMS-W on a wide scale. The Climax Molybdenum Company, a Division of American Metal Climax, Inc., located at Climax, Colorado, was among the first companies in the Americas in recent years to support the development of HIMS-W to recover weakly magnetic minerals on a large scale. The Climax Molybdenum Company, one of the world's largest underground mines, processes about 40,000 tons per day of ore, from which about fifty million pounds of molybdenum disulfide are produced each year. By-products include tungsten, iron pyrite, tin, and some monazite.

Jan 1, 1968

By understanding the ore breakage characteristics and utilising comminution circuit modelling, the benefits of SAG mill feed size manipulation can be readily assessed. This paper looks at the benefits and problems of total secondary crushing of mill feed and advocates partial secondary crushing as a means of improving grinding efficiency for certain hard ores. Case studies are reviewed for new and existing SAG mill circuits.

Jan 1, 2007

By Michel Dagbert

With the growing demand for energy, explora¬tion for uranium deposits has dramatically increased in the past few years. Since the early ages of uranium exploration, natural gamma ray logs has proved fast and convenient in locating and evaluating mineralized bodies. However, natural gamma ray logging is an indirect measure of the uranium content and in some circonstances, the radiometric signal can be 'altered in a very significant way. To check the presence of any distortion, statistical comparision of radiometric counts and chemical assay values for the same samples is generally performed. Of course in such studies, one inevitably mixes samples where the radiometric signal is really proportional to the uranium content and samples where this signal is influenced by many other sources or distorted before being recorded. Nevertheless, the average tendency leads to correction factors which are then applied to every radiometric reading in the deposit or portion of the deposit where the statistical comparison is made. To account for the "average" character of these correction factors, a parameter of uncertainty like a standard deviation of possible errors is associated to each corrected value. This parameter can be latter used in an ore reserve estimation when the weights assigned to samples should depend on the precision of their grade. The purpose of this paper is to present a new regression technique which can be used without any restrictive hypothesis like a normal or lognormal distribution. Different standard ways which can be used to statistically compare radiometric and chemical assay values for the same sample and to derive a correction factor for the latter will first be reviewed, after a general discussion of disequilibrium problems. The discussion and the examples will be restricted to the case of natural gamma ray logging in drill holes but the methods presented in this paper can also be applied to data from the radiometric probing of a working face or a truck load in an operating mine.

Jan 1, 1979

By S. K. Kawatra, J. A. Halt

"Low-grade iron ores are finely ground and processed to recover iron-bearing minerals. However, the iron-rich powder is too fine to be used directly in blast furnaces and most direct reduction furnaces, and must be agglomerated or pelletized into a usable product called “pellets.” Bentonite clay has traditionally been used as a binder to make iron ore concentrate pellets, but it is expensive and contaminates the pellets. Organic binders may be a viable alternative to bentonite clay. This manuscript reviews organic materials that have been used to make iron ore concentrate pellets. Included are discussions on the development of organic binders, common pros and cons to organic binder use, organic binder structure and its effects on agglomeration, and additives to organic binders.IntroductionIron ore agglomeration turns finely disseminated, low-grade iron ores into a usable product, called pellets. During pellet production, fine-grained, iron-rich minerals are liberated from low-grade ores, concentrated, and agglomerated into round balls that are thermally hardened or cold-bonded to form pellets. Pellets are then transported to blast furnaces or direct reduction operations where they are used to produce iron; iron is further refined and shaped into the wide variety of steel-based industrial and consumer products in use today.Over the years, pellet production has become increasingly more prevalent with the increased exploitation of low-grade iron ores and increased iron and steel consumption around the world. In 2010, world iron ore pellet production was 388 Mt. It could be argued that the current pellet production process has been enabled by the use of binders – practical experience has shown that it would be nearly impossible to agglomerate iron ore concentrates using rotary drums and discs without using a binder."

Jan 1, 2015

By R. R. Klimpel

The use of chemical reagents in flotation has been studied in great detail, e.g. (1, 2, 3, 4). Important progress has been made on the understanding of surface chemistry concepts involved in flotation. However, such scientific studies are often very complex even for rather simple idealized systems. Therefore, the deve- lopment of predictive or quantitative tools for plant engineering evaluation is still in the evolutionary stages. Until further progress in fundamental work is achieved, the operating engineer is faced with semi-empirical physical testing of reagent changes on several sizes of the actual equipment under study to be confident of the economic viability of any change.

Jan 1, 1980

By F. G. Crack

This paper is based on the philosophy that the management structure and strategy required for mine construction and development projects are different from that required for routine production and maintenance operations in a mine. The purpose of this paper is to present an integrated approach to the management of mine development projects. This paper highlights the four major aspects of the integrated approach. First, a brief review of the various types of contract documents is provided. Second, the management organization of a project is discussed. Third, the project manager's role and the field supervisor's responsibility are briefly outlined. Finally, the use of scientific techniques for planning, scheduling, executing, and controlling projects is presented. Each of these aspects is important; however, each by itself cannot be expected to bring order or unifying impact to the complexities involved in overall project management. When all these aspects are coordinated, the result will be a sore successful project from conception to completion.

Jan 1, 1982

By S. Boone, J. N. Shirlaw

INTRODUCTION Soft, near normally consolidated clays generally provide ideal conditions for tunneling using Earth Pressure Balance (EPB) shields. It is easy to turn the clay into a low permeability, viscous spoil that can be pressurized to maintain the required face pres-sure. Despite this, there are records of relatively large, and often quite variable, surface settlements over EPB driven tunnels in soft clay. To understand this variation it is necessary to identify the various factors that contribute to settlement. Tunnels in soft clay are ideal for study these effects in the field, as an unsupported void will close almost immediately, due to the low shear strength of the soil. In the short term, deformation in a soft clay will occur at constant volume, so short term surface settlements can be related to movements at tunnel level. In this paper, data will be used from EPB driven tunnels in Singapore marine clay. In all of the cases discussed the lining consisted of precast concreted segmental rings erected in the tail of the shield. The marine clay in Singapore has been studied, among others, by Tanaka et al.(2001). The clay typically has a plasticity index of 40 to 60, and the undrained shear strength (SU ) can be derived from: SU = 0.22 svo' where svo' is the current vertical effective stress. Figure 1 shows a summary of undrained strength/depth plots used for design on two sections of Singapore’s North East Line. The stability of a tunnel in clay can be assessed by comparing the stability number (N) of the tunnel with the stability number at total collapse (NTC). The stability number can be calculated from the equation: N = (?Z + q – Pt)/SU where: ? = the average unit weight of the ground above the springline of the tunnel Z = the depth to the springline of the tunnel from the ground surface q = the surcharge Pt = the support pressure within the tunnel The stability number at total collapse depends on the size, geometry and depth of the tunnel, and typically ranges between 5 and 9. Charts for assessing the number at total collapse are provided in Kimura and Mair (1981), based on the testing of model tunnels in a geotechnical centrifuge. For tunnels in the Singapore marine clay, the stability number of an unsupported tunnel would be over nine if the tunnel ois more than 8m deep. As a result, a support pressure

Jan 1, 2007

By W. Pariseau, S. Puri, S. Schmelter

Introduction Joints often have profound effects on elastic properties and strength of rock masses and therefore on rock slope stability. In surface mine slopes, joints are usually too numerous to be taken into account individually, so an equivalent properties approach is necessary. Previous work that treats a jointed rock mass as an equivalent composite material of joints and intact rock between resulted in excellent agreement between estimated and true rock mass elastic moduli in case of fully persistent joints (e.g., Pariseau, 1999) Rock mass failure mechanisms on joints and through intact rock were retained obviating the need for an equivalent rock mass strength. Impersistent joints, which are not continuous on a joint plane, can be accommodated (Pariseau, 2003). However, slope stability analysis of a large, deep copper mine indicated a need for computational efficiency even in two-dimensional analysis of vertical sections through pit walls (Puri, 2006). The analyses by the popular finite element method used small elements of bench size near the pit slopes and much larger elements away from the pit walls. The reasons for the graded meshes were numerical quality (small elements) and computational economy (large elements). Small elements contained only a few joints, while the larger elements contained hundreds of joints and led to impractical computation run times (days!). A new modeling procedure that recognizes sufficiently large elements as representative volume elements (REV’s) overcomes this obstacle. A companion improvement recognizes the concept of a representative area element (RAE) in association with joint impersistence and the probability that joint material occupies a considered site. Both are steps towards greater model realism and a practical slope stability analysis procedure. Consider the open pit mine photo shown in Figure 1(a), a companion plan view in Figure 1(b), and the problem of slope stability where high slope angles favor economics but low slope angles favor safety. These competing factors in surface mining generally result in some sections of the mine that have unstable slopes that require monitoring and perhaps mitigating action to improve stability. Although stability of a given section may not be easy to quantify with a single number, useful design guidance can be obtained from a finite element analysis of a proposed mining sequence leading to future pit slopes and profiles. The presence of water from rain and snow melt indicates a need for a coupled analysis that links water pressure, effective stress and total stress (Schmelter, 2001).

Jan 1, 2007

By Alex G. Sciulli

According to recent estimates, nearly 30 percent of the material from underground coal mining in the United States is rejected on the surface as waste. This accumulation of almost three billion tons of coarse and fine coal refuse is primarily from coal cleaning operations. The marketable coal is intermixed with sandstone, clays, shales, and carbonaceous rocks, which must be removed and disposed of in an acceptable manner. The disposal of mineral wastes is a major environmental concern. The most common types of environmental problems associated with coal refuse disposal are acid formation, erosion, and sediment control. Other areas of concern include air and water quality, combustion control, mine sealing, and reclamation. This paper focuses on these topics and attempts to outline special environmental features which can be incorporated into the design and construction of a coal refuse facility. Often these measures can minimize and possibly eliminate the potential environ- mental hazards. It is imperative, however, that these measures be adopted prior to and during construction, because of the relative ease and cost savings compared to post-construction implementation.

Jan 1, 1983

By James R. Evans

Orange County, adjacent to the populous and industrialized Los Angeles area, is the fastest growing county in California and is second only to Los Angeles County in total population (figure 1). In addition to an equable climate and a burgeoning economy, there is room in Orange County for growth and expansion. Growth over the past decade has resulted in rapid development of industrial tracts, housing areas, and shopping and commercial centers, mostly in former agricultural areas. The value of new building in 1972 was a record amount-over I billion dollars (figure 2). Problems relating to the proper direction of growth are the concern of both government and the private sector. The intelligent and efficient use of areas to be developed requires planning for the conservation and protection of the County's natural beauty and for the orderly development of natural resources. The availability of sand and gravel, which is used in many phases of construction, will be a major controlling factor in the future of the County.

Jan 1, 1974

By D. F. Lillie

The milling plant of Bicroft Uranium Mines Limited has been in operation for a period of 2 years. During this time many changes, both physical and chemical, have been made in an effort to improve plant efficiency. Investigational work has been facilitated greatly by the fact that the Bicroft mill consists to a large extent of two parallel circuits; however, the number of inter-related variables is great, and it is difficult to measure, let alone control, all of them simultaneously. The Bicroft mill staff feel that much progress has been made, but there is much yet to be done and many questions still to be answered. The property of Bicroft Uranium Mines is located in Cardiff Township in the Highlands of Haliburton County, southeastern Ontario, some 140 miles northeast of the city of Toronto and 70 miles due north of Lake Ontario. The ore-bearing dikes are classed as mafic, non-segregated, syenite and granite pegmatites. Microline, microline perthite, quartz, uraninite, uranotharite, fluorite, titanite, cyrolite, calcite, pyrite, pyrrhotite and molybdenite as important accessories. The mill was designed for a nominal capacity of 1,000 tpd with all sections except grinding, leaching and clarification capable of 1,500 tons. Feed was entered to one mill unit on October 29, 1956 and to the second parallel line on November 13. The total feed rate was at 1,000 tpd on November 15. The first precipitate was produced on November 12. By August 1958 the average entry had been increased to 1,410 tpd without the installation of any additional major equipment. The overall recovery has been greater than 93,0 pet do several months, but a recovery of 92.5 pct appears to be most economical. Operating time for 8 months in 1958 has been 97.1 pet although all grinding mills were relined during this period.

Jan 1, 1959