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By D. W. Rierson

Svedala has a ported kiln process, designed specifically for the direct reduction of pelletized or natural lump iron ores. The process is called ACCAR. This ported kiln technology has been coupled with Svedala's GRATE-KILN System, for iron ore pelletizing, into the GRATE-CAR System for the reduction of fine grained ore concentrates through the steps of agglomeration, induration and reduction in a single production line. Additional applications of the GRATE-CAR System include: the recovery, agglomeration and reduction of steelmill waste oxide fines and the use of minus 6mm size natural lump ores to produce a highly metallized pellet agglomerate. Svedala conducts tests to confirm certain operating parameters through process simulations of the GRATE-CAR process. These tests are conducted at Svedala's Process Research and Test Center, Oak Creek, Wisconsin. Svedala also assesses issues of process variables, heat transfer, equipment sizing, retention time and endothermic demand associated with iron ore reduction using natural gas or coal as the reductant. The purpose of this paper is to present the GRATE-CAR processing concepts, the commercial production line and the testing procedures associated with investigating potential raw materials.

Jan 1, 1997

By Ademar Lopes, Victor Silva, Heitor Silva, Gustavo Usero, Roberto Rolo, Lucas Pereira, Gabriel Moreira, João Lague, Arthur Endlein, Matheus Feitosa

Grade control models are supposed to provide a higher resolution than long-term or interim models. However, available grids from diamond or reverse circulation drilling do not provide that required resolution. In this study, a geostatistical workflow is proposed to integrate available grade data from production blastholes with its operational parameters taken when drilling downhole. The step of geological logging can be replaced by machine learning models for classifying blastholes into lithologies and hardness classes. Next, indicator and ordinary kriging use the classified blastholes and previous data for estimating the model domains and the grades within each one. The workflow is applied in a world class iron ore mine in Brazil.

Jun 25, 2023

By S. Davis

It has been accepted almost as dogma that sampling contributes most of the error to fluctuations in gold fire assays. Differences in assay results on pulp material from the same sample interval are frequently regarded as evidence of the presence of the so called "nugget effect" or relatively coarse gold particles. It is true that if coarse gold particles are present they contribute to substantial sampling fluctuations, but while the process of sampling is probably the major source, the analytical process cannot -be completely ignored as a possible contributor to erratic assay results. In order to maintain a stable assay process, the analytic part of the system must be kept in control as well as the sampling and preparation part. One way of monitoring the performance of the analytic system is to systematically assay standard material whose sampling characteristics are carefully controlled. Gold assay standards are not prepared nor can they be prepared to account for both sampling and analytical errors. It is not possible to send coarse material to a lab for preparation (which includes both the processes of comminution and splitting) followed by fire assay, to make inferences only about the fire assay process. Because most gold ores are very heterogeneous, sampling errors would in most cases completely mask the contribution of the analytical errors. Assay standard material is prepared only to assess the accuracy and variability in the fire assay process.

Jan 1, 1995

By Ingo Kaundinya

"INTRODUCTION Tunnels are important elements of the German federal road network which have to cope with an increasing traffic load especially of heavy goods vehicles. In order to ensure a high availability and safety level for these key infrastructures, big efforts are required by owners and operators in order to keep all infrastructures up to date. This mainly relates to current European regulations regarding tunnel safety (Directive 2004/54/EG) [1] and their national implementation, which must be fulfilled by many existing tunnels. In this context the German Federal Government spends more than 1.1 billion € for the structural and operational retrofitting of more than 120 existing road tunnels. As many of these projects have recently been finished or are still under progress, experiences and best practices could be reported. After the description of current requirements in national standards regarding structural fire protection, the paper gives an example for an upgrade of a tunnel’s fire protection by installing an additional fire protection system, which is a quite costly measure. Another possibility to fulfill current requirements regarding structural fire protection could be the numerical proof of sufficient bearing capacity of a structure. The influence of different fire loads and fire durations on the bearing capacity of typical road tunnel structures have been investigated in a research project funded by the German Federal Ministry of Transport (BMVI) and coordinated by the Federal Highway Research Institute (BASt). The results of this research project have been used to define a procedure how to investigate existing tunnels regarding structural fire protection. Finally, the research results have also been the basis for a current update of national standards for tunnel construction. REQUIREMENTS REGARDING STRUCTURAL FIRE PROTECTION The structural fire protection design of road tunnels is usually based on temperature-time curves of exposure to fire, for simplicity called ""fire curves"". The fire development after fire growth is illustrated with fire curves. Fire curves can be subdivided into a heating-up phase, a fully developed fire phase, and if applicable, into a cooling-off phase (cf. Figure 1). During the fully developed fire, the fire remains at a certain temperature level."

Jan 1, 2016

By T. D. Barkand

A recent coal mine elevator accident resulted in serious injuries to four passengers. The miners struck the ceiling of the elevator car as it collided into the overhead structure at an estimated speed four times faster than normal. In response to mine hoisting accidents, safety authorities have required additional protective devices on new and existing mine hoists and elevators. This paper examines supplemental safety circuits and new devices designed to provide ascending car overspeed protection on mine hoists and elevators. These devices provide needed protection that eliminates the hazard or significantly reduces the potential of a mine hoisting accident.

Jan 1, 1995

By Stephen R. Holsinger

During the year of 1972 ASARCO's Silver Bell embarked upon a project that would test a modified grinding circuit using direct digital compute against the performance of its other five grinding lines, implementation and subsequent outcome of that project has been previously reported in a paper by Pitts, et al. The main purpose of the present paper will be to describe major problems incurred during the implementation of this control strategy on the remaining five primary grinding lines and their solutions. Also included will be a discussion of changes made in operating philosophy and methods used since that report was written. Description The Silver Bell concentrator is a 10,436 tonne/day copper sulfide plant supplied by ore from two separate open pits approximately 4 km. apart. As well as receiving ore from these two main pits, ore is received from several exten¬sions of both pits. The ore bodies are variable in rock type and the extensions have increased this variability. Grinding is carried out by six primary, single-stage, 3.20M x 3.66M Allis Chalmers grate discharge ball mills in closed circuit with four Krebs D20B cyclones (Fig. 1). Each of these grinding circuits is controlled by four independent but interactive control elements. These control elements are the fresh ore feed, the water added to the feed end of the mill, the water added to the ball mill discharge sump, and the cyclone feed pump speed. At the present time, instrumentation for measuring primary process variables consists of Merrick weightometers for fresh ore feed, orifice plates coupled with differential pressure transmitters (dp's) for incoming water measurements, back-pressure diaphragm coupled with a dp for estimating ball mill discharge sump level, a magnetic flowmeter for determining cyclone feed velocity and a nuclear density gauge for measuring cyclone feed density. All measuring instruments develop electrical signals that are measured as 2-10 V DC "inputs o the computer interface.

Jan 1, 1977

By A. K. Sinha

Ever since electricity was introduced, electrical injuries and electrocutions have continued to occur regularly in coal mines. An analysis of the two hundred cases of electrocutions which occurred during 1956 to 1970 indicated that over 50 percent of the incidents were caused by the ineffectiveness or lack of electrical grounding. Research shows that various types of arbitrarily designed ground beds are in use and that the ground resistance and current are often very high and subject to wide seasonal variations. The results of a recent research are discussed in this paper.

Jan 1, 1980

By J. Gathje

Normal process development for pressure preoxidation treatment of refractory gold ores starts with laboratory batch autoclave tests, followed by operation of a continuous autoclave pilot plant for confirmation of the various process options. This scheme will generally provide good information showing the effects of such variables as grind size, time, and temperature, and will indicate whether autoclave preoxidation will produce solids that are amenable to cyanidation for gold recovery. However, several recent test programs conducted over the last three to four years have shown that the cyanide leaching responses of treated solids from the batch tests can differ considerably from those achieved by the continuous pilot plant autoclave. Examples are provided to demonstrate these differences and to show the importance of extending the laboratory investigations to include the use of a semicontinuous autoclave test procedure. A test case is cited in which this procedure provided data that could not be obtained from batch tests.

Jan 1, 1995

By George W. Toth

In a recently completed research project for the U.S. Bureau of Mines - Twin Cities Research Center, a computerized costing procedure for uranium in situ leach mining was developed by the NUS Corporation. This costing procedure, termed a cost model, employs a process engineering approach for estimating total project costs as well as equipment and manpower requirements for uranium in situ leaching operations in either Texas or Wyoming. 80th capital and operating costs are generated by the model along with the overall production cost/pound U3O8 subject to a known rate of return. Conversely, rate of return on equity can be solved for, subject to a given price of yellowcake. During the course of the research effort, this cost model was applied 50 typical site conditions in Texas and Wyoming. Sensitivity rests were conducted for numerous key parameters to determine the cost influence attributable to incremental changes In parameter values. The results of these sensitivity tests along with a description of the in situ leaching cost model have been documented as Bureau of Mines, Mineral Research Contract Report No. J019912, "Cost and Sensitivities Analysis for Uranium In Situ Leach Mining," dated March 1981.

Jan 1, 1981

By George C. Hawley

Historically, mica usage has been low because of limited availability; as a by-product of many small processors, who have lacked motivation for extensive R & D. The discovery of a 30 million ton deposit of 85-90% pure phlogopite, when existing markets are only 300,000 tonnes annually, forced the search for large volume new markets. Increasing costs of energy, petrochemicals, plastics and glass fibers, and environmental constraints on asbestos usage, have led to new opportunities for mica as a reinforcing filler. Extensive research and developments; in beneficiating, processing and treating mica, has been necessary to satisfy these new markets.

Jan 1, 1981

By C. J. Im

As market conditions become more competitive, U.S. coal producers are faced with the increasing need to find methods of lowering production costs in their operations. One area worthy of investigation is the additional recovery of fine coal that is normally discarded in the preparation plant waste stream. United Coal Company is currently assessing the capability of the Deister Flotaire, a new application of a column flotation cell, to recover additional, saleable coal from preparation plant waste. This paper presents the technical data obtained to date. Work with a 0.76 m (30 inch) diameter cell installed in a plant shows considerable coal recovery is possible from a nominal 0.595 run (28 mesh) X 0 refuse stream.

Jan 1, 1986

By A. J. Kroha, N. Wesis

Most copper mines produce both ore and low-grade "leach" rock or acid waters that contain recoverable copper. The sulfide ores pre¬dominate, and a portion that is too low grade for milling to produce concentrates for smelting, but has to be mined and trucked away anyhow, may be leached successfully with acid in dumps. Most of this leach material consists of sulfides and silicates or carbonates, and if the gangue is such that it consumes a high quantity of acid, this factor may rule out a leach operation. There are also valuable deposits that contain mostly acid-soluble copper, or occasional sulfide ores from which a sulfide concentrate can be roasted and acid-leached to produce a copper-bearing solution. Finally, there are milling ores in which the lesser part of the copper is acid-soluble and can be precipitated with iron or synthetic inorganic precipitants that produce metallic copper or copper sulfides that will float with the sulfides. Ordinarily, ores that contain copper associated with the sulfur ion, such as in the minerals chalcopyrite, chalcocite, bornite (and others), are milled to produce a 25-30% Cu concentrate for smelting, while a lesser amount of acid-soluble copper may be converted from solution to cement copper on iron scrap. A fast-growing percentage of such copper, however, is removed from solution with exchange resins or organic compounds in organic carriers such as kerosene (solvent extraction), then eluted with strong acid and subjected to electrolytic precipitation either in marketable form or as anodes that can be refined further. From the point of view of conventional copper smelting, copper flotation concentrates and cement copper are of chief interest in this chapter. Table I is a condensed open schedule for concentrates that generally run between 25 and 35% copper, and much less frequently as low as 12-15% or as high as 65-75% copper, the former being due to intimate relationship with pyrite (like the former United Verde Extension), and the latter representing such ores as the Bolivian Coro¬coro ore in which the copper is in the form of chalcocite in sandstone. These extremes are no longer common. When they occur, a special purchase schedule has to be negotiated. Included in Table 1, copper precipitates (cement copper) generally run from 70-85%a copper, and the same basic purchase schedule is used as with flotation concentrates. Sulfide Flotation Concentrates The sulfide copper concentrate produced in the mill as a flotation froth, with water then added for transportation of the heavy mineral particles from the flotation cells to thickeners, may run 60-80% water by weight, and the removal of water down to 25-50% by weight by means of thickeners, followed by further dewatering by continuous vacuum filters to 7-18% moisture by weight (depending on size of solids by screen analysis and also by content of clay) is a critical operation. Mill operators would like to produce a filter cake with 7-9% moisture content, but even with the help of steam on the filter this desirable condition is seldom realized when the concentrate is as fine as 80% -325 mesh. More commonly, the final concentrate is reground in pro- to produce best copper recovery and grade of concentrate (or molybdenite separation). In those cases, increasingly frequent, the filter product may not be a cake at all, but a mud that is hard to handle-even requiring a thermal dryer. Greater difficulty of form¬ing a manageable cake often comes from the copper-molybdenum separation by flotation, because the alkaline sulfides and hydrosulfides, or cyanide, or other similar reagents used for the separation, may leave the now relatively molybdenite-free copper concentrate even more difficult to filter. Handling a wet filter cake is difficult enough when its destination is only a short distance away-a matter of yards rather than miles. In those cases the filter cake may be thermally dried near the point of production, using rotary or multiple hearth, or fluidized-bed dryers. Alternatively, the concentrate may be pumped or carried in slurry form to the smelter and filtered there, or it may be spray-dried and compacted. For transportation to a smelter just a few miles to a few thousand miles away by ship or railroad other factors may be important, such as: in shipping by sea, avoidance of spontaneous combustion; in shipping by rail, losses by leakage if too wet or by wind and sun if too dry. It is the responsibility of the millman-usually the mill superinten¬dent-to make sure that his concentrates are in satisfactory condition when they leave the mill so that they meet these requirements: 1) They must have been accurately sampled and dry-weighed, the latter meaning that a moisture determination and gross weight must have been taken. 2) They must be dried sufficiently when necessary to prepare them for safe transportation. 3) They must arrive at the smelter with reasonable likelihood that they can be check-weighed and sampled fairly and equitably,

Jan 1, 1985

By Miguel Carrera, Jose Arnal, Julian Ortiz, Maarten Haest, Fouad Faraj

ShovelSense is a complete solution for bulk ore sorting at the mine face that combines robust shovel bucket mounted X-ray fluorescence sensors, with real-time material classification, fleet management integration and supporting cloud-based data infrastructure, to ensure each truck always goes to the right destination. Correct truck destinations are a challenge in many current mining settings because the material classification of a truck, which is the key input into the truck destination, is often based on estimates at block resolution (~20x20x15m), with limited adjustment for blast movement and mixing. In this contribution, we present a geostatistical, data driven approach to estimate the potential value of bulk ore sorting at the mine-face for seven open pit mines. The approach is premised on the assumption that truck loads within blocks follow a simple gamma distribution. Standard ore control data from the seven mines have been used to estimate the potential bulk sorting value at these operations and ShovelSense-integrated dispatch datasets have been used to validate these estimates for specific benches on which ShovelSense operated. ShovelSense results validate the ore control data-based value predictions and show how at mine-face bulk ore sorting value varies between distinct deposit types. In general, deposit types that have higher bulk ore grade contrasts, display more potential ShovelSense value. The manuscript concludes with the limitations and opportunities of applying the proposed value estimation approach to evaluate at mine-face bulk ore sorting opportunities for open pit operations.

Jun 25, 2023

By R. Keith Evans

Potential large scale uses in secondary battery systems and fusion reactors together with a steady increase in demand for conventional applications, particularly in aluminum smelting, has resulted in an increasing interest in lithium and its compounds. Current world requirements are met from both granitic pegmatites and non-pegmatitic sources and although the emphasis in this short paper is on the pegmatitic minerals, and, in particular, on their "direct usage" applications, a short summary of other sources and requirements for chemical production is included for the sake of completeness.

Jan 1, 1977

By T. W. Mattioli

Restrictions on fugitive emissions and crushing operations continue to get tighter. Although fugitive dust control in the gold mining industry has historically been difficult, there are now several options that can make dust control both effective and efficient. Using new technology, mine operators can now design and maintain a dust control program that can reduce product losses, decrease equipment failure, cut cleaning costs, and meet air quality compliance standards as well as provide a better working environment. This paper discusses the sources of fugitive emissions related to the gold mining and crushing operations and describes effective new control methods.

Jan 1, 1998

By Daniel J. Dobbels, Eric R. Drooff, James R. Wheeler

Construction of Boston’s C11A1 Section of the Central Artery (I-93)/Tunnel(I-90) Project beneath the existing MBTA Red Line Subway at South Station required excavation to depths in excess of 33.5 m (110 ft) through permeable sand and gravel subject to almost 27.4 m (90 ft) of hydrostatic groundwater pressure. As tacked drift tunnel underpinning system was designed to provide positive support to the active station prior to mainline tunnel construction. As part of this system, a permeation grouting program was developed to facilitate tunnel excavation. This paper describes the design and implementation of the grouting program, in sit testing, and the overall performance of the grouted soil mass.

Jan 1, 1999

By I. R. Miller

The purpose of this discussion is to try to offer some suggestions as to how the zinc and lead markets will perform over the balance of the decade, review some of the factors influencing production and come to some conclusions regarding prices. The past 2 years have been simply awful in terms of consumption, prices and overall profitability for the industry. In looking ahead to the year 2,000, it is probably not a bad idea to look back about 20 years to see where we have been and how we got to where we are. Zinc consumption from 1950 to the early 1970's increased steadily and rather nicely at an annual average rate of close to 4.0%. It was starting in the early 1970's that things started to get volatile and at times, difficult for the industry. The general economic boom of the early 70's was marked by annual consumption rates close to 5.0 million tons, a level not seen again for over 10 years. The First Oil shock of 1973-74 was followed by the sharpest consumption drop in memory. In two years, Western World zinc consumption fell 27%. While our chart shows a rather dramatic recovery thereafter, the average annual growth rate since 1970 has only been a rather modest 1.5%.

Jan 1, 1992

By Jonathan M. Pachter

The Bailey mining complex is located in Greene County, PA in the valley of Talley Run, a small stream tributary to the Enlow Fork of Wheeling Creek as shown on the location map insert found on Figure 1. The Bailey complex includes two Pittsburgh No. 8 Seam deep mines (Bailey and Bailey #2), a coal preparation plant, and two refuse areas (for disposal of coarse and fine processing waste). Clean coal is shipped from the complex via a 24.15 km (15 mile) rail spur, which connects to the Monongahela Railroad in Waynesburg, PA. In early 1981, Consolidation Coal Company (Consol) began a permitting effort aimed at beginning construction of Bailey, the largest underground coal mining complex to be built to date in the United States. Construction of the Bailey Mine began in December, 1981, less than six (6) months following the first application submittal. Clean coal was produced by the Bailey Preparation Plant in September, 1984. Effective project planning, including early formulation of a project planning team, identification of "red flag" issues, innovative responses to environmental questions and development of environmentally sound project designs, all contributed to the relatively quick permitting at Bailey.

Jan 1, 1985

By S. Tuylu, D. Adiguzel, A. Bascetin, E. G. Savas, O. Ozdemir

"Paste Technology has some advantages comparing to other surface disposal methods. With this method, it is possible to pump high concentrated solid materials to very long distances without segregation problem. However, it is known from Literature that pumpability of materials is completely related to rheological properties of the materials. For this reason, in this study, the rheological parameters (e.g. yield stress and viscosity values) of the paste material prepared from Pb-Zn tailings at various solid ratios (60-85%) were obtained in order to determine the pumpability of the paste material. Additionally, slump tests, which are usually used in mining industry, was carried out. Finally, regression analysis was made between the slump values and rheological parameters of the paste material to find a correlation between two independent parameters. The results showed that there is a great correlation between the slump values and the yield stress of the material. On the other hand, there was no correlation with viscosity: ty=430.28-16.877Sv (R2= 0,94; ty: Yield Stress (Pa), Sv: Slump Value (cm)). Based on the results obtained from this study, it can be clearly concluded that the slump values obtained from a simple test can be used to predict yield stress values of paste materials. INTRODUCTION Recently, surface disposal of mining tailings has become one of the biggest problems in mining industry in terms of environmental issues. Therefore, several researches have been focused on this subject in order to find a best solution or method for this problem. These studies showed that paste technology has some advantages compared to other conventional methods. Tailings thickened to the extent show no segregation of particles during transportation, and relatively small amount of settling occur post-deposition which are often called “paste” (Cincilla et al. 1997). Paste tailing technology has two applications in the industrial area as underground paste fill and surface paste disposal. Paste tailing technology was first started to be used in Grund mine, Germany in the late 1970s (Grice, 1998). Process tailing material, water and - if needed - cement materials, which are the components of the paste, must be blended to obtain a homogenous mixture suitable for pumpability. Consistency behavior of the paste material is determined using a conventional sedimentation cone concrete test (slump). Meanwhile, the slump value of the paste must be between 200 and 250 mm for surface applications (Newman et al., 2001). Additionally, in the case of underground paste system, the optimum slump value must be between 150 and 200 mm. On the other hand, the water content and density of paste at certain slump value change depending on the amount of fine particles in the tailing sample. While moisture of the material will be high, its density will be low because of surface area of particles which increases with decrease in particle size (Meggyes and Debreczeni, 2006). For pumpability of the paste to long distances, slump value of the paste must be good enough to be accepted for use of surface paste disposal. Paste material has a yield stress because they show non-Newtonian behavior and generally is classified as a Bingham plastic fluid. (Cooke, 2008; Theriault, 2003). In addition, the appropriate viscosity value of the paste along with the yield stress must be between 100 and 250 Pa (Bascetin et al., 2013; Bascetin et al., 2014; Bascetin et al., 2015)."

Jan 1, 2016

By Tamás Megyeri, Dániel Borbély, Vera Szánto, László Kovács

"Upon completion, the Bátaapáti deep underground radioactive waste repository complex (National Radioactive Waste Repository - NRWR) will provide safe final disposal for low and medium-level radioactive waste from the Paks Nuclear Power Plant in Hungary. The waste emplacement is in progress and one of the main concerns related to the development is ensuring a feasible plan is in place for permanent repository closure. Detailed hydrogeological observations and safety assessments revealed that the clay filled fault zones penetrated with the access tunnels play determinant role in the long-term confinement performance of NRWR, so they need to be plugged. Detailed mechanical analysis and numerical modelling of the tunnel constructed through a fault zone is presented here to gain a better understanding of the rock mass and rock support behavior. The effect of stress reflection is demonstrated and its consequences are described in detail. A recommendation is given for tunnel support design in such conditions. INTRODUCTION Crossing fault zones is a challenging task during tunnel construction and can lead to failure if not properly accounted for during design and construction. These failures are can be caused by high water inflow or stability loss of the rock mass surrounding of the tunnel (examples can be found in CEDD, 2015). The prediction of the magnitude and development of displacement has a considerable impact on issues during tunnel construction. Changes of rock mass stiffness or structure ahead of the face, to a great extent influencing stresses in the vicinity of the tunnel and thus deformations (Steindorfer, 1998). Further research with numerical simulations was carried out, and the qualitative and quantitative evaluation of displacement vector is presented in (Grossauer et al, 2005). The concern associated with crossing fault zones was considered during the design of the Bátaapáti deep underground radioactive waste repository complex (NRWR) in Hungary. In the case of NRWR project, the fault zones have an impact on maintaining the stability of the surrounding rock mass as well as the final sealing of the repository. As part of the final closure design, a complex numerical modelling program was completed to gain a better understanding of the mechanical behavior of the rock mass and the monitoring observations during construction. Even though the need for the presented study is closely related to the NRWR project, the results can be used in any project related to fault zone crossing. The methodology used to better understand the impact of crossing fault zones for the NRWR project is presented in this paper. First the project background is introduced including an overview of the underground facilities, the aim of the repository, and the construction experience gained from the first crossings of the fault zones. After that the in-situ and laboratory testing is summarized. The numerical modelling methodology is detailed in the next section. Not only the outcome of the model is presented, but the results are compared with the experience and observations made on site. Since a good agreement is found between the actual and the modelled behavior, the conclusions are based on the validated result. To fulfill the aim of the paper, the stress reflection concept is introduced to explain the mechanical behavior of the fault zone and the rock support; furthermore straightforward suggestions are made for rock support design."

Jan 1, 2016