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By Barbery G., A. Maachar, M. Bourassa

Laboratory testing for flotation circuit design is carried out on a routine basis in a large number of lab- oratories throughout the world. Usually a sample of ore, assumed to be characteristic of the material to be processed in a future plant, having a mass of the order of a few kilograms, is ground in a batch ball or rod mill, in the presence of water and some re- agents (for pH control, etc. .. ); after dilution to the adequate solid concentration, the resulting slurry is transferred into a laboratory batch flotation machine, having a volume which can vary from 0.5 to 8 10 -3m3. Various types of machines are in common use, manufactured by the same companies that market industrial size flotation machines. The operation of flotation in the laboratory is carried out by an experienced technician, who uses parameters such as impeller rotation speed, air flowrate addition, froth height, froth removal rate, in conjonction with reagent addition sequence, to "test" the ore. After flotation, the froth products ("concentrates"), and the solids remaining in the cell are filtered, dried, and analyzed for the various elements of interest. For flotation circuit design, a kinetic test is usually carried out, in which the froth products obtained at various times are collected to assess the time response of the ore in the laboratory machine with the combination of reagents used. Although this description is rather sketchy, it enables to outline some of the major difficulties in using laboratory data for flotation circuit design: 1. ore sample representatively 2. feed preparation procedure (laboratory compared to industrial scale) 3. laboratory procedures (type of ma- chine, type of operation, experimental errors. . . ) 4. flotation machine scale-up 5. flotation kinetic scale-up 6. flotation circuit design using laboratory data These various points will be discussed in the present paper, in order to stress the problem areas as well as possible solutions. ORE SAMPLE REPRESENTATIVITY It has been stated many times that one of the major problem in designing a flotation circuit is the representatively of the sample to be taken. Sampling of an ore body during prospection and evaluation is a specialized field, which has been progress in with the introduction of geostatistics. The application

Jan 1, 1986

By R. W. Barbaro, R. V. Ramani, K. V. K. Prasad, P. T. Luckie

Discussion by A. Unal Barbaro et al. (1990) implemented a tedious study aimed at delineating the short-scale spatial variability of sulfur in a coal seam. It is not possible, however, to extend their conclusions to any other coal seam or use their in any other fashion, because the background geology was not presented. Nor were the conclusions accompanied by geological interpretations in the paper. In addition, an unfortunate printing error occurred where the total sulfur (determined by High Temperature Combustion) variogram in Fig.3 was duplicated in Fig. 2 instead of the seam height variogram. A more serious error, however, has been committed in the definition of the variable, seam height. What is defined as seam height by Barbaro et al. (1990) is, in essence, mining height, and is not a regionalized variable that should be studied by geostatistical methods directly. Despite the fact that the variable, seam height, is not defined in the paper explicitly, the following quotation discloses the fallacy: "All roof rock that exceeded 1.8 m (6 ft) from the floor was not taken because the longwall was operated to not mine more than 1.8 in (6 ft) unless the coal seam height exceeded 1.8 m (6 ft)." From this definition, the seam height, and therefore the sampling height, is equal to the actual coal thickness, if the coal thickness is greater than 1.8 m (6 ft). Otherwise, it is equal to the sum of the coal thickness plus the thickness of the roof rock that complements the thickness to 1.8 m (6 ft). In the latter case, the seam height is constant and equal to 1.8 m (6 ft). It is possible to conduct a variogram study on a pool of samples that are realizations of two different variables. But the conclusions derived would not belong to any one of the two variables uniquely and, therefore, do not possess any significance. Geostatistical analysis is irrelevant for the sum of multiple regionalized variables formed by arbitrary selections. In a two-seam setting, for example, the mining height, as defined by the thickness of one seam at one location and the thickness of both seams at another location (due to quality and/or minimum thickness considerations perhaps), should not be used in the calculation of one common variogram. The two seams should be modeled separately. They can then be combined according to the specific purposes of the study. On the other hand, if a constant is added to a regionalized variable (to incorporate dilution perhaps), the variogram of the new variable will not change. Barbaro et al. (1990), surprisingly, does not give any geological interpretation of their results despite the fact that most of them can be explained by the origins of sulfur in a coal seam. The presentation of the results of a geostatistical study with no reference to the geology of the deposit is uninformative. It may also be misleading for the potential users of geostatistics, It is not unusual to find nuggets of pyrite in coal seams. In such cases, pyritic sulfur will probably display a spatial structure for only a very small distance that will appear in the experimental variogram as no spatial correlation. This very well known phenomenon is called the pure nugget effect in geostatistics (Journel and Huijbregts, 1978) and perhaps can explain the lack of correlation found for pyritic sulfur content. The lack of correlation found for the total sulfur content may also be explained in the same way because the total sulfur content is dominated by the pyritic sulfur content in this case study. One should notice, however, that the situation may completely be reversed after cleaning the coal. Not all of the inorganic sulfur should be expected to be in the form of pyrite nuggets in a coal seam. It may also be disseminated in the coal seam and it is expected that it follows a certain spatial structure. However, an existing spatial structure may be masked by including a part of the roof rock, rich in sulfides, into the seam thickness in an arbitrary fashion because areas having a sandstone roof sometimes are known to show a higher sulfur content due to the downward percolation of solutions rich in iron sulfides (Clark, 1979). Plants use sulfur in their growth processes. Much of this sulfur is bound organically during peat accumulation and coal formation (Cecil et al., 1978.) This suggests a spatial structure of some sort for the organic sulfur. However, it is not possible to test this hypothesis because the results obtained by the authors for the organic sulfur content are not given in this paper. For this reason, the conclusion that simple average of the nearby samples would provide the best unbiased estimates is questionable for organic sulfur and is not based on any substantial supporting evidence. It is suspected that no spatial structure was detected and this was due to high sampling and laboratory analysis errors. Before concluding that sulfur variability in the seam at the location of study was random, a more detailed study for the disseminated non-pyritic sulfur should have been conducted, not for the sake of scientific curiosity only, but also due to its utmost importance with regard to coal cleaning and emission control. Pyritic sulfur can be cleaned to a considerable extent, whereas organic sulfur can not, making the emission control strategies highly dependent on the spatial distribution of the organic sulfur (Knudsen, 1981). In the light of these facts, one wonders why Barbaro et al. (1990) did not present the results of their study for the sulfate and organic sulfur content. References Barbaro, R.W.. et al., 1990, "Short-Scale Spatial Variability of Sulfur in a Coal Seam,' Mining Engineering, Vol. 42, No. 11, pp. 1267-1269. Cecil, C.B., et al., 1978, "Geology of Ccontaminants in coal," report prepared for Environmental Protection Agency, North Carolina, 123 pp. Clark, W.J.. 1979, "An interfluve model of the upper Freeport coal Bed in part of western Pennsylvania," unpublished MS thesis, University of South Carolina, 57 pp. Journel, A.G., and [Huijbregls], Ch. J., 1978, Mining Geostatistics, Academic Press, London, 600 pp. Knudsen, H.P., 1981. "Development of a Conditional simulation model of a coal deposit," unpublished PhD dissertation, The University of Arizona, 109 pp.

Jan 1, 1992

By R. Vogt, H. C. Mult, F. L. Wilke

INTRODUCTION At present, the capability of production planning software based on Linear Programming (LP) is still limited to the optimization of the single LP-run. This is due to the LP-model itself which cannot consider the interdependencies between individual LP- runs. With regard to planning of selective mining this limited way of optimization frequently leads to situations, where the remaining and accessible ore blocks do no longer allow to produce ROM-ore in the qualitative composition required by the ore processing plant. However, many of the aspects taken into consideration when setting up production plans built from mutually dependent LP-runs cannot be modelled in a system of linear equations. They are thus unsuited for treatment with LP and have to be taken care of by the planning engineer without any assistance by the system. The KBS currently under development is intended to assist the planning engineer in designing a production plan under special consideration of the combination of consecutive LP-runs and blending beds. It is not necessarily intended to find the optimum solution within a given planning situation which is, anyway, hard to determine due to the multitude of influences. The objective is rather to work out a good and - from the practical point of view - feasible production plan. The new aspect with respect to mine planning is the integration of expert knowledge and experiences via the KBS into the planning process in order to support the planning engineer. The planning system is being developed in close cooperation with an iron-ore open pit mine. COMPONENTS OF THE PLANNING SYSTEM The software is being developed on a workstation under UNIX and comprises the components LP, CAD-module and the KBS. The applied multi-goal LP-algorithm is an in-house development of the Department of Mining Engineering at Technical University Berlin. It was already successfully implemented within other mine planning programmes and was only slightly adapted to the specific needs of the present system. Within individual LP- runs it finds the optimum qualitative composition of ore production in the sense of the selected optimizing criterion and under the given restrictions: i.e. it determines tonnages to be mined from blocks in order to optimally meet the requirements of the ore pro- cessing plant. A CAD-module based on the commercial SURPAC package in combination with a simulation device is used to graphically depict the block model and progress of mining. Both LP-algorithm and CAD-package are integrated in the KBS. It has been decided to use the shell NEXPERT OBJECT as it is a hybrid system which supports both rule-based and object-oriented knowledge representation. MINE-MODEL AND LP-MODEL KBS have to be tailor-made for specific planning problems. Therefore, it had to be decided which specifications of the iron-ore mine should be represented in the model. As the limited possibilities of a university institute do not allow to develop a KBS for mine planning which is ready to use in industry, it was decided to concentrate on those characteristics that can be regarded as typical for iron-ore surface mines and that seemed to be suited for treatment with knowledge-based techniques. The following chapter summarizes the most important features of the mine model. The description of the requirements to the mine's sales products is followed by an outline of the applied LP-model. Mine model • The model of the mine as it is used for planning consists of • the block model of the deposit, • the mobile equipment, • stockpiles and blending bed and • the requirements to the sales products. The deposit is described by a block model which contains data on the chemical composition, LOI, grain size and tonnages. Grain size was included as it is important for the two sales products of the mine. Furthermore, it is known whichs blocks require and which don't require blasting; this is relevant to the assignment of loading equipment to individual blocks. The blocks are devided in three categories: • ore, which will directly be taken to the blending bed; • waste, which will be put on the waste dump; and • blocks which will be either transported to the blending bed, to stockpiles or to the waste dump depending on the specific planning situation. This decision is made during planning. Neighboring blocks are combined in mining areas to which the loading equipment is individually assigned. Mobile equipment comprises shovels and wheel-loaders as well as trucks. The characteristics of the loading equipment are important for their ability to load different blocks and for the permissible degree of their re-positioning etc. The mine disposes of a blending bed for homogenization of the production, of a waste dump, and of several stockpiles with different ore qualities. The requirement to make only limited use of the stockpiles for economic reasons is included in the KBS. According to long term planning two commercial products have to be produced, which differ both in grain size and qualitative composition (TABLE 1). Their mass-proportions in the blending bed have to be within a fixed range. Not considered in long term planning is the occasional need for lump ore, which occurs at very short notice and has to be produced in a "campaign-like" manner. This requires the total re-arrangement of all plans for on- coming blending beds and would therefore be ideally suited for

Jan 1, 1996

By Houston Kempton, Tim Dyhr, Andy Davis, Mark Bloom, Connie Travers, Dave Atkins, Jeff Deen

INTRODUCTION Historical mining activities in the Robinson Mining District, in the Egan Range of east-central Nevada (White Pine County), resulted in five open pits (Veteran, Tripp, Liberty, Ruth, and Kimbley Pits), several large mine-waste dumps, and numerous underground workings developed for copper, zinc, gold, lead-silver, and manganese (Figure 1). Groundwater percolating downward through the copper sulfide ores produced a blanket-like supergene zone enriched in copper, which has been one of the primary objectives of mining since the early part of this century. To obtain permits for future operation, an EIS was conducted that evaluated future pit- lake chemistry, groundwater quantity and quality issues, transport of water and solutes subjacent to a proposed tailings impoundment, and the potential for waste dump acidification. AREA HYDROGEOLOGY The mine area is on a ridgeline draining to the Steptoe Valley to the east and to the White River Valley to the West. Precipitation in the Robinson Mining District ranges from a low of 15 cm (6 in.) per year in the valley to more than 50 cm (20 in.) per year in the surrounding mountains. Evapotranspiration exceeds precipitation, leading to a net loss of water from the basin as either surface flow or underflow to the north at Currie, Nevada (Eakin et al. 1967). The White River Valley, which lies to the south and west of the mine area, includes approximately 4,200 km2 (1,620 mi2) of the upper part of the Colorado River Basin. This valley is 112 km (70 mi) long, 32-48 km (20-30 mi) wide, and like Steptoe Valley, evaporation and transpiration exceed precipitation (Maxey and Eakin 1949). The local groundwater flow system in the Robinson Mining District and the proposed mine area is dominated by two main components: (1) a groundwater mound or high between the Ruth Pit and Murry Springs, and (2) a shallow groundwater flow divide that follows the ridge crest south of Ruth, which contains the Veteran and Tripp pits. Besides the alluvium, the main hydrostratigraphic units controlling groundwater flow are limestones, shales, and sandstones, along with the intrusive quartz monzonite stocks. Based on the piezometric surface measured in October 1993, three distinct groundwater regions were identified. In the Giroux Wash area, an apparent fault runs between WCC-G1 and WCC-

Jan 1, 1995

The transfer of Langmuir-Blodgett (LB) monolayers of collector molecules to the surface of mineral crystals is an excellent method for the study of collector adsorption phenomena and is of particular significance in the understanding and development of improved flotation separations. The fatty acid/fluorite and amine/sylvite potation systems were investigated by the Fourier transform infrared/internal reflection spectroscopy (FT-IR/IRS) examination of transferred LB monolayers of collector molecules at the surfaces of the mineral crystals. The [ir]-A isotherm study of fatty acid and amine, typical anionic and cationic collectors, showed the different packing trends dependent upon the pH and/or type of subphase. This result clearly reflects the effect of pH on compound formation and dissociation of the collector molecules. Identification of the collector adsorption state and confirmation of the validity of the FT-IR/IRS adsorption density equation were included in this research effort. This allowed for a detailed discussion of collector adsorption phenomena. As a result of these experiments, it is possible to relate the hydrophobic character of a surface to the nature of the adsorbed collector structure and thus describe flotation behavior in greater detail. For example, it appears that in the case of oleate/ fluorite the equivalent of an LB monolayer can form spontaneously during adsorption from solution, and this structure is known as a self-assembled monolayer (SAM). Such a state is significantly more hydrophobic than the oleate structure which forms on the calcite surface at an equivalent adsorption density.

Jan 1, 1996

By D. D. Carda, S. Dunn, W. E. Dunn, T. Storbeck

Gold chlorination which will be presented in this paper is limited entirely to the use of gaseous chlorine in a fluidized bed reactor at elevated temperatures in the absence of aqueous water. The only water used in this process is that which is used in scrubbing the effluents. Low temperatures are defined as those below 500°C. All experimentation for this paper was carried out on a laboratory scale, with all calculations extrapolated frpm this. Some of the data is drawn from work with other projects which have been operated at pilot and larger scale.

Jan 1, 1998

By Samuel B. Romberger

During the formation of most epithermal precious metal deposits geochemical changes can be characterized by the introduction of large amounts of silica, minor amounts of sulfur, and trace quantities of gold, silver, and other metals. During the same process, large amounts of alkalis and alkaline earth elements, such as sodium and calcium, and sometimes aluminum are removed from the rocks. In deposits hosted by carbonate sedimentary rocks large amounts of carbonate may be removed before or during the silicification process. Small amounts of iron may be introduced by the mineralizing solutions, however, in most epithermal deposits this may be a relatively immobile component and is locally derived. The results of these changes are deposits consisting of up to 90 to 95 percent silica as quartz veins, breccias, silicified host rocks, and jasperoids containing a few percent sulfides, mostly pyrite, and a few parts per million gold and silver. Other minerals commonly but not always present are barite, arsenic, antimony, or mercury sulfides, and tellurides. Often gold shows a spatial, and therefore an apparent genetic, relationship to pyrite in primary ores. Alteration and gangue assemblages sometimes include sericite and/or adularia, and parts of the deposits may contain alunite and/or clay minerals. Even though the value of these deposits usually lies in their gold content, silver is usually present in larger quantities than gold (Romberger, 1990). Based on these general characteristics observed in a large number of epithermal deposits, together with processes occurring in active geothermal areas to which they are often genetically linked (Henley, 1985), a model is proposed whereby deposition occurs along a subhorizontal zone separating environments of contrasting chemistries. The upper environment is characterized by oxygenated acid waters and the deeper environment by a reducing near neutral aqueous system. The model describes the chemical changes occurring as a result of mixing and/or boiling within this zone. It builds upon the physical models proposed by Romberger (1988a) for disseminated gold deposits and Bonham (1988; 1989) for volcanic rockhosted deposits.

Jan 1, 1990

By J. Bleeck, G. Mahrt

INTRODUCTION Many excellent papers have been written and presented regarding agricultural lime and the subject has been covered extensively in books concerning soil science, fertilizing, limestone, etc. We felt challenged to construct a paper that wouldn’t merely be a repetition of past presentations. Therefore, we chose to begin with factors having the most impact on plant growth realizing that the audience is likely to be relatively unfamiliar with this subject. So, we decided to dig deep into the roots of soil and plant growth science in order to better understand the significant role of limestone in this important area. The soil with the higher plants growing upon it forms an open and dynamic ecosystem in which there is flux of energy and material (heat, light, water, nutrients, mineral and organic matter, organisms). Plants root in the soil and then take up oxygen for root respiration, yielding energy for nutrient uptake, water for turgescence and transpiration and use in photosynthesis and the mineral nutrients needed for the synthesis of organic matter. Above the ground the plant takes in light energy and heat using them in photosynthesis to assimilate atmospheric CO2. Photosynthesis is the basic life process of creation. However, we will look at the root room where soil, water, air, and heat economy and the supply of nutrients determine soil fertility and the important role lime plays here. Location is the primary point to be considered when discussing agriculture (climate and soil); secondly, the dynamic and mobility of the minerals (plant nutrients) in the soil; and thirdly the conversion of organic substances into humus. And then there is a great deal of water involved in the growing process. In order to produce 1 kg of dry plant substance, 300 to 800 kg of water must be pumped through the plant and is then lost by transpiration. Cereals, vegetables, and sugar beets consume between 400 to 500 kg of water during the growing process and long-growing crops, such as alfalfa and clover, can consume 600 to 800 kg of water per kg of finished product. Without water, a plant cannot grow. (Reference 1)

Jan 1, 2006

By Joseph Fernandez, Luminita Calin

INTRODUCTION The Tunnel and Reservoir Plan (TARP), commonly known as the Deep Tunnel Project, is the daring and unique concept developed by the Metropolitan Water Reclamation District of Greater Chicago (MWRDGC) as a two-phase solution to the problem of waterway pollution and basement flooding due to sewer system backups created by the combined sewer over-flows (CSOs). Phase I, primarily for pollution control, includes the tunnels, shafts, connecting sewers and structures. These Phase I facilities intercept, during wet weather events, the CSOs within the MWRDGC TARP service area. The tunnels and shafts provide temporary storage of the CSOs for subsequent treatment at Water Reclamation Plants (WRP) before discharging into waterways. Ultimately, upon completion of Phase II, the tunnels, located from 45 m(150 ft) to 100 m (330 ft) below ground, will convey the CSOs to reservoirs for storage until WRP capacity is available. Phase II, which includes the reservoirs, pumping facilities, additional tunnels and shafts, has been planned to serve a dual purpose. The reservoirs provide storage of excess flows during wet weather events, those beyond the WRP capacity. Following the wet weather events, the stored flows can be conveyed, at controlled rates, to the WRPs for full treatment. These excess flows previously would have resulted in both overflows into the waterways and eventually backups in the local sewer systems. The storage capacity of the reservoirs also will promote the full utilization of the conveyance capacity of the tunnels and shafts, which in turn results in optimizing the conveyance capacity of the existing sewer systems. Therefore pollution control is further enhanced by allowing for capture of a greater volume of CSOs and greatly minimizing backups in the local sewer systems. The TARP project consists of four systems: O’Hare, Mainstream, Des Plaines, and Calumet, covering a total area of 971 km2 (375 mi2) and serving the City of Chicago and 51neighboring communities as shown in Figure1. With only two contracts left before completion of the Phase I tunnels, the 12.7km (7.9mi) Torrence Avenue leg, to be awarded in 1997, and the 12.4km (7.7mi) Little Calumet leg unfunded as of yet, both part of the Calumet System, the pollution control phase of the project is now 86% complete (Table 1).

Jan 1, 1997

By Joseph R. Hatch

Previous studies of the Thompson and Temperly ore bodies show alteration aureoles in the Middle Ordovician host-rock limestones, dolomites and shales extending laterally as much as 60m from the ore. Thickness, texture, carbonate and clay mineralogy, major-, minor-, and trace-element contents; and carbonate carbon and carbonate oxygen Isotope compositions of the host rocks all have been significantly altered by hydrothermal processes. Organic geochemical analyses of samples from the basal 8-cm-thick carbonaceous shale bed and immediately overlying limestone bed of the Quimbys Mill Member of the Platteville Formation, collected from within the ore, the alteration aureole, and the unaltered country rock show: 1) solution of carbonates and residual concentration of organic matter, 2) decreases in the volatile (Si) and pyrolytic (S2) hydrocarbon yields In samples of the basal shale from within the ore, 3) decreases in the relative amounts of the lower molecular weight (<nC25) saturated hydrocarbons in samples from the ore, and 4) no changes in either the organic carbon isotope composition or the Rock-Eval or biomarker thermal maturation Indicators. In light of the 125-170°C depositional temperatures documented for the early marmatitic sphalerltes in these deposits by fluid Inclusion studies, the lack of thermal alteration of the organic matter suggests a relatively short mineralization episode, a conclusion supported by the relatively simple sphalerite band stratigraphy. The decreases in the volatile and pyrolytic hydrocarbon yields for shale samples from the ore supports a model for mineral deposition which obtains sulfide through reaction of brine sulfate and indigenous organic matter.

Jan 1, 2013

By John L. Bolles

The mineral processing engineer must have some knowledge of mineral identification, and the greater his knowledge in this field, the greater will be his potential to do his own specialty well. Most metal-bearing ores and industrial minerals require processing of some kind to produce a salable product or a material suitable for further processing. The types of treatment are many, but each one is aimed at separating a mineral or family of minerals from others. Obviously, then, it is basic to all mineral processing to identify the minerals present in the raw material and in intermediate and final products resulting from processing. Very often the mineral processing engineer will enter planning while exploration of a mineral prospect is going on. He will examine hand specimens, drill cores and cuttings, bulk samples, and materials in place. He must identify the valuable minerals, their occurrence and associations with other minerals, and the identity of diluents or host rock. An exact and detailed knowledge of the raw material will usually suggest the treatment processes that should be investigated in the development of the most suitable scheme for that particular raw material in that particular location. Physical properties of a mineral often lead directly to a process for separating that mineral from others. Some well-known examples of these are: magnetic separations for such minerals as magnetite, rutile, ilmenite; gravity separations for gold, silver, galena. cassiterite; flotation separation for most sulfides; distillation of mercury from cinnabar; and selective dissolution in specific solvents such as halite or sylvite in water, gold and silver in cyanide solutions, copper oxides, carbonates, sulfates and silicates in dilute sulfuric acid. These processes and many others are described in detail elsewhere in the Hand¬book and so it is not appropriate to give details here, except to emphasize that processing is aided greatly by complete mineral identi¬fication. Developing a treatment scheme for a new mineral discovery is an exciting assignment fag the mineral processing engineer, but much more of the usual work relates to quality control in existing plants. Here too, mineral identification is important. Some processing is aimed toward producing an end product which will meet exacting specifica¬tions regarding purity or allowable diluents. In this situation it is vital that contaminants be identified so that the process can be adjusted

Jan 1, 1985

By D. R. Baughman

Copper oxide deposits that may be considered for in situ leaching exist in unique geologic settings, different from uranium deposits. A large number of interrelated factors must be considered in wellfield design. Many of these factors are hard to quantify and are directly related to geologic features of the orebody. This paper discusses some of the important factors in wellfield design and suggests an outline for a design procedure.

Jan 1, 1988

By Larry Whittle

Gold milling activities generate waste streams containing dissolved metals and cyanide, among other contaminants, that require some form of treatment prior to discharging to the environment. To date, this treatment has involved the destruction of cyanide through natural degradation, chemical oxidation, or a combination of both followed by the precipitation of metal hydroxides. Although these processes have been proven to be capable of achieving. regulatory levels for metals and cyanide, they often generate significant levels of cyanate and, at times, ammonia, that may be toxic to rainbow trout and thus undesirable for regulatory compliance. Tallon Metal Technologies Inc. (Tallon) has developed Vitrokele™, a chelating resin that has the ability to capture metals (including gold), both total and weak acid dissociable cyanide, as well as thiocyanate from both slurries and solutions. The loaded resin is eluted and returned to the recovery circuit while metals and cyanide are recovered from the eluent stream. This process was demonstrated at pilot scale at B.P. Resources' Hope Brook Gold Mine in Newfoundland, and more recently at Canamax's Bell Creek Mine near Timmins, Ontario. The results from these pilot programs showed metal and cyanide recoveries of 99% or more with the concentration of cyanide and metals in the effluent reduced to levels with minimal or no toxicity to rainbow trout. Furthermore, neither cyanate nor ammonia were generated from the process. The loaded resin was effectively stripped and re-introduced to the recovery circuit; cyanide was captured in the eluent stream and made available for recycling to the mill leaching circuit; and metals were captured for re-use, sale or disposal. This paper summarizes the results obtained from the pilot tests at the Bell Creek and Hope Brook Gold mine sites.

Jan 1, 1992

By K. Soucek

Extraction of hard coal reserves, but also of mineral resources in general in the underground coal mines includes in itself many anthropogenic interventions in the earth crust. These interventions are closely associated with the solution of stability problems in the mine and with occupational safety. One of potential applications of such geo-technical interventions in rock mass is the chemical grouting technology. The chemical grouts are very suitable to affect the physical and chemical properties of rock mass. With the present development of chemical grouts, primarily based on polymers, there come into existence new composite materials during the grouting process (the so called geo-composites), which properties significantly differ from the original rock. An important factor, which affects a choice of chemical grouts and at the same time the effectiveness of grouting technology is the resulting properties of geo-composites. The properties of a coal geo-composite differ from the ones of the unmodified coal seam. On the basis of the laboratory geo-mechanical tests, the differences of properties of both material types were found. While the pure coal samples were destroyed very brittle and explosive, the geo-composite samples were making higher level of tenacity and load bearing ability after overstep of material strength. The presented paper discusses the different behaviour of these original and modified coal samples. The X-ray CT scans were used for observation and demonstration of failure evolution in these samples of pure coal and coal geo-composites after compressive loading.

Jan 1, 2006

By James R. Arnold, John G. Mansanti

BACKGROUND Gold Fields Mining Corporation started their Chimney Creek operation in late 1987. The dump leaching operations commenced in October and the mill startup took place during the last half of November. The milling process at Chimney Creek is almost a conventional CIP plant with the major exception being the installation of a second thickener as a washing stage prior to CIP12 &apos; A simplified flow sheet is shown in FIGURE 1. Most of the gold in the mill is liberated by grinding in cyanide. After grinding in a conventional SAGball sill circuit, the cyclone overflow reports to the thickeners. The thickener overflow from the first thickener is pumped through 5 stages of carbon columns for gold adsorption. Slurry from the thickener underflow is leached, washed, re-thickened and pumped to CIP. After approximately 6 hours of retention time in CIP, the slurry is pumped to tailings for final deposition. The carbon from CIP is transferred to the third stage of carbon adsorption and then advanced as normal through the columns and strip circuit. Loaded carbon is stripped at approximately 1500C (3000 F) and 55 kPa (80 psi). After 4-5 hours of stripping, the grade of four tons of carbon is reduced from 5,140-6,860 g/t (150¬200 oz/ton) to typically 17 g/t (0.5 oz/ton). Zinc precipitation is used to recover the gold from the pregnant solution. After a water wash, the stripped carbon is transferred to acid wash or to the kiln feed tank for reactivation. All strip lots are acid washed in 3.0% HC1 and all stripped carbon passes through the kiln for thermal reactivation. Then the carbon is quenched, sized and stored in tanks or bins for transferral to CIP or to the carbon columns. Loaded carbon from the dump leach circuit is treated in a circuit consisting of five carbon columns which are parallel to the mill circuit. No distinction is made between the mill or dump leach carbon once the gold has been stripped from them. CARBON FINES HANDLING Once the carbon has been transferred from the carbon columns, every effort is made to minimize carbon attrition and to recover the carbon fines that have been generated in the process. Wemco Hidrostall pumps are used for transferring all carbon in

Jan 1, 1990

In April 2012, work started on one of the most ambitious construction projects of Austria?s ÖBB-Infrastruktur AG ? the New Semmering Base Tunnel. Scheduled for completion in 2024, the 27.3 km (17 mile) tunnel through the Semmering range will cut the journey time on the south railway connecting Lower Austria and Styria. Before con-struction can begin at the entrance portal and tunnel bores, the work concentrates on the above ground sections with the clearance and excavation of the entrance ramp site. The slope is currently being excavated by the Baubeteiligungsge-sellschaft m.b.H. BBG-GRAZ. The cut face is shored up with shotcrete and permanent self-drilling anchors produced by Minova MAI, an economical bolting solution for softer rock formations. After completing the surface construction work, the twin-bore, single-track tunnel will be cut through the Semmering by conventional and continuous excavation beginning in 2015. Once completed, the New Semmering Base Tunnel will be one of the longest rail tunnels in Austria. It will act as a vital link on the route crossing Europe?s heartland from Northern Germany to Italy, which currently has to climb the steep Semmering pass between Gloggnitz and Mürzzuschlag on a historic mountain railway recognized as one of the UNESCO World Heritage sites. The old route imposes considerable limitations on heavy freight trans-port, and passenger trains have to slow down to a speed of around 50 kph (31 mph). The New Semmering Base Tunnel will bypass this bottleneck and cut 30 minutes off the journey time between Vienna and Graz. It represents a particular advantage for freight traffic, since the new route?s minimal incline of 8.4 percent removes all obstacles for even heavy freight trains.

Jan 1, 2013

By S. Orr

Flow in heaps and dumps is a two-phase flow phenomenon, though for many applications it could be simplified as unsaturated flow. Unlike saturated flow (typical of ground water flow), where permeability is independent of other hydraulic parameters, unsaturated flow permeability depends on the degree of saturation and/or on capillary pressure. Several recent field studies suggest that flows in heaps and dumps tend to concentrate in preferred pathways, bypassing much of the ore. Different preferential flow phenomena are triggered, promoted and influenced by different heap structures, by pretreatment, by the composition of the leach solution and by application rates and schedules. The structure of heaps and dumps is determined and affected by every stage of their construction -from blasting to crushing to conveying and stacking of the material. History of heap construction and operation combined with monitoring, mapping and certain field tests can reveal patterns of flow and transport in a given heap and can shed light on the physical mechanisms behind it. Such insight can provide a first cut on heap leaching enhancement by either eliminating or bypassing physical barriers to flow and leaching. Based on the understanding of the heap structure, advanced flow and transport models can be constructed and further enhance the insight. This insight is essential for understanding cause and effect and can provide a basis for decision making. Further, based on the physically-based models, an intelligent control agent, such as MRDS1, can learn and build a self-learning representation (or a model) of the complex heap leaching system, followed by dynamic optimization and planning of heap leaching and heap rinsing.

Jan 1, 2003

By A. Redfern, B. G. D. Smart

The ultimate objective of the development work presented in the paper is to devise an appropriate method for relating powered support specification to geological conditions and mining practice, thus enabling the longwall method of coalmining to be successfully implemented throughout the varied coalfields of the world. The paper begins with a review of various methods previously used to determine support rating, each method being analysed with regard to its source, the logic and theory which constitute the method, and subsequent recommendations. An alternative approach is then proposed, the Powered supports being considered to be one of three foundations, sharing load with coal ahead of the faceline which acts as another foundation, and the waste behind the supports which acts as the remaining foundation. The load/deflection characteristics of the coal and waste, which in turn are related to geological conditions and mining practice, are shown to affect the load transfer on to and hence the specification of the powered support. In particular the characteristic of the waste foundation is shown to be affected by the caving mechanism, an alternative form to that mechanism accepted for British conditions being proposed and designated "parting-plane-controlled" caving. A finite element solution for a beam on elastic foundations was used to translate the above concepts into numerical results, these results being compared with data from Indian and Chinese operating conditions. The Indian case study examines the effect of using a conventional chock type of support in a one-web back mode when mining towards a roof break generated ahead of the face, while the Chinese case study compares measured subsidence over a longwall with "blocky" roof conditions with the model results. The comparisons are encouraging, suggesting that the approach used is worthy of further development.

Jan 1, 1986

By Bernhard C. Koch, Richard W. Hutchinson, Bernhard Free

The Tertiary Silverton caldera collapsed asymmetrically to the southwest 27.5 m.y. ago. Resurgent volcanic activity is expressed by a quartz-monzonite stock at the southern caldera margin and by the Red Mountain porphyry stock in the northwest of the caldera which were emplaced at 25 m.y. and 11 m.y. b.p., respectively. The apical Eureka graben formed between 27.5-22.5 m.y. ago and transects the Silverton caldera southwestward. In the northeast it terminates at the 22.5 m.y. Lake City caldera margin. This graben as well as caldera faults provide dominant structural control for base and precious metal deposits. The epithermal gold mineralization of the Gold King mine is proximal to the Red Mountain stock in the north central part of the caldera. Here, west-northwest striking faults related to the collapse of the caldera intersect northeast faults of the Eureka graben. Superimposed upon earlier base metal mineralization at 23 to 22.5 m.y. b.p., the 11 m.y. younger gold metallization was a distinct event at Gold King. Supergene enrichment of the hypogene precious metal assemblage resulted in a metal tenor of up to 3,016 g/t gold and 1,088 g/t silver. The deposit is hosted by quartz-latite tuffs, flows, and flow breccias of the Burns Formation. Veins, up to 5 m wide follow steeply southeast and northwest dipping structures of the Eureka graben. Subsidiary veins up to 2 m wide branch off the main lode and are termed "flat veins" because of their conformable strike but shallow dip of 45&apos; to 50&apos; toward the northwest. Meteoric water-dominated hydrothermal solutions generated by the Red Mountain stock ascended through intensely fractured, ferruginous

Jan 1, 1988

By H Sohn, R. R. Fernández

Microscopic gold inclusions in sulfide minerals cannot typically be extracted by conventional cyanidation. The gold ore must be pretreated to oxidize sulfide minerals associated with gold and to remove other minerals and chemical compounds that interfere in the cyanidation leaching process. The recovery of gold from refractory sulfidic-carbonaceous ores after roasting pretreatment was studied. With the addition of an oxygen-containing gas (e.g., air or oxygen-enriched air), sulfides of iron, zinc. copper and other metals are tramformed into solid oxides or sulfates and gaseous sulfur dioxide at elevated temperatures (e.g., 400° to 800°C). When the sulfides are oxidized, individual particles lose internal volume and become porous, allowing leaching solutions access to the gold. The structural characteristics of the calcine have a strong influence on the eventual recovery of gold by cyanidation. These characteristics are strongly dependent on the roasting conditions and roasting method. The optimum combination of particle size, roasting temperature, reaction time and oxygen concentration in the off-gas was determined for a typical refractory sulfidic-carbonaceous ore from the Carlin Trend in Nevada. The tests were run using a 150-mm laboratory rotating tube furnace and 160- and 200-mm.fluid-bed pilot roasters. Laboratory results indicated that the maximum gold extraction was 92%, with the head and the leach residue containing 5.7 and 0.4 g/t (0.166 and 0.013 oz/st) Au, respectively. The maximum extraction was achieved under the following roasting conditions: a top particle size of 150 yin, a residence time of 2 hrs, a roasting temperature of 475°C and with 40% oxygen (by volume) in the off-gas. Similar results were obtained from circulating fluid bed (CFB) pilot-plant testing. The maximum gold extraction was 90%, with the head and leach residue containing of 5.1 and 0.5 g/t (0. 149 and 0.015 oz/st) Au, respective) v. These results were achieved with the same particle size and oxygen concentration as in the laboratory testing but with the roasting temperature of 500 ° to 525°C and a residence time of 30 min.

Jan 1, 2000