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By D. A. Wilson

A laboratory method for processing lead battery scrap has been developed that eliminates virtually all of the S02 emission that normally is produced when a reverberatory or blast furnace is used. The method employs a hydrometallurgical step to convert the PbS04 to PbO and CaS04 by mixing the battery scrap with a slurry of Ca(OH)2. The reaction is exothermic and goes to completion in 2 to 5 minutes at room temperature. The dried product is mixed with a KC1-NaC1 flux and carbon and reduced at 6500 C. A minimum charge ratio of 3 parts of carbon to 100 parts of battery scrap is required for reduction. The process produces lead reductions and recoveries as high as 97.0 and 94.9 percent, respectively. The KCI-NaCl flux is recovered using a standard counter-current water leach followed by evaporation.

Jan 1, 1976

Environmental leaching methods such as the toxicity characteristic leaching procedure (TCLP) test or the ASLP test are used to assess wastes including mining and mineral processing by-products. These tests were initially designed as regulatory tests simulating conditions in a co-disposal municipal landfill waste facility and do not allow leaching performance to be predicted in site specific scenarios. Further, the assumptions underlying these tests tend to define worst-case scenario leaching conditions for by-products from the mineral industry disposed of outside a co-disposal scenario.   Recently some characterisation leaching test methods have been issued as ISO/TS standards (ISO/TS 21268-3:2007 and ISO/TS 21268-4:2007) that aim to define the leaching performance of soil type wastes and to predict the leaching performance of these wastes over long time frames. In this preliminary work these standards are trialled for by-products from the mineral industry and the quality of information that can be obtained from them compared to TCLP/ASLP tests.   Where clear comparisons were possible, the TCLP/ASLP tests were typically found to extract a similar, or often greater, amount of component than (i) tests conducted at the same pH but without the acetate buffer or (ii) for test conditions aimed at simulating long-term leaching in an eluate defined by the waste type. While further work is required, the preliminary results indicated that the tests may provide reliable long-term estimates of leaching performance of wastes in soil types of variable pH and provide a suitable basis for long-term planning for optimal waste handling, treatment or even re-use.

Jun 1, 2010

By J F. Costa, A Grigorieff

Coal mining reclamation is a worldwide concern. This paper presents a Brazilian example where the economic aspects of reclamation are considered. The basic goal of the study is to transform the reclamation procedures into an economic process by integrating the use of land after reclamation and developing profitable post operations. Some of the activities considered in this case study include: using of the coal pit as a landfill, considering appropriate landfill design and construction; using the area for forest development; and developing grassland, forests of native species and the construction of ponds for fishery. This study will conclude that the mined area is: capable of storing large quantities of waste at a competitive cost and at a low environmental risk; commercial forest development is economically feasible; and the lakes and other reclaimed areas can be integrated into the landscape creating a fresh water supply and an area to raise cattle.

Jan 1, 2005

By Robert E. Kirby

The potash mine and refinery of AMAX Chemical Corporation is located 26 miles east of Carlsbad on the north end of the potash producing areas. This property was developed between 1949-1952 and production of muriate began in August 1952. Two shafts, located 400 feet apart, were sunk to the ore zone 900 feet below the surface. A 15-foot diameter ore hoisting shaft equipped with a 1000-ton ore pocket, crusher and automatic skip loading equipment plus a 20-foot diameter service shaft with a curtain wall for exhausting air were provided. Sylvinite production to date has been from the first ore zone, stratigraphically, the lowest of ten identified mineralized zones in the salt section. This first ore zone has been the top producer of Muriate of Potash over the years. In our area we do have some very marginal mineralization in the third ore zone some 35 feet above the first. We have not mined from this seam as yet but we are investigating the economics very closely with access slopes and an experimental mining panel. Our ore seam is generally flat lying but does have local gentle rolling with grades from 2% to 5%. The mine was designed and developed from the beginning with conventional mining practices and rail haulage from the mining section to the ore pocket. As mining progressed in the early years, production increased from less than one million tons annually to a maximum of 3,935,000 tons in 1964. Efficiency also kept pace with the increasing production and continually improved until the efficiency of the production crews reached an average of 100.3 tons per manshift in 1964.

Jan 1, 1974

By I. Iwasaki

Platinum group elements (PGE) are often found in arsenide and sulfarsenide minerals as inclusions in copper-nickel bearing gabbroic rocks. Major losses of platinum in flotation have been attributed to the poor recoveries of the arsenic minerals. Flotation characteristics of a synthetic nickel arsenide and arsenopyrite were studied through rest potential and oxidation rate determinations, surface analyses, and micro-and batch-flotation tests. The use of nitrogen during grinding and flotation together with copper activation improved the PGE recoveries from Duluth Complex ores. The flotation behaviors of the sulfarsenide minerals are still unresolved. The role that electrochemical interactions play among arsenides, sulfarsenides, sulfides and grinding media on their f10atabilities is discussed.

Jan 1, 1988

By J. C. Gilbert

"IntroductionTHE PURPOSE of this paper is to discuss in a general way the heavy oil situation in Western Canada, and particularly to acquaint those in the oil industry who have not participated in the heavy oil end of the business with past and current developments. The author has covered •briefly the location of the heavy •and asphalt-base oil fields, their geology, current drilling and production practice, the location of the •refineries capable of handling heavy crude, their products and markets, and the problems and future prospects of the industry.Location and Description of FieldsThe term heavy oil is generally used to denote crude of 20.9° gravity and under. Since the products from heavy oil are also to a lesser extent obtained from asphalt-base crudes of higher gravity, some mention will be made here of the latter; but in general the emphasis of the paper is on heavy oil as such.In the order of their production to date, the heavy oil fields and those light oil fields which produce asphalt-base crudes are : Lloydminster, Taber, Vermilion, Conrad, Princess, and Wainwright.The Lloydminster oil field has not only produced most of the heavy oil, but also is currently the largest producer, supplying at the present time about 80 per cent of the heavy and asphalt-base crudes produced in Western Canada."

Jan 1, 1952

By A McGann, P J. Coghill, D Miljak

"It is widely accepted that bulk ore sorting could provide a step change in processing of low-grade ores. For this to be achieved a rapid, accurate bulk sensor of grade is required as well as sufficient natural variation of grade in the ore stream. CSIRO trialled a magnetic resonance (MR) sensor capable of measuring chalcopyrite, and by inference total copper grade, on the 1300 t/h Ridgeway portal conveyor. The gauge produced an unprecedented set of data down to a 1 t scale for months at a time. This paper analyses the measured variation and compares it to a simple model derived from the bore core data and mixing within the block cave mine.The model of the small-scale (<1000 t) variation is developed using fractal methods which have successfully been used in the past for copper porphyry deposits showing a power law or near power law distribution of copper grade and vein sizes. These fractal distributions have two features highly favourable for segregating ore, a large proportion of low-grade pods and the large-scale spatial clustering of grade. We show the presence of fractal ordering in bore core data at short range in the Cadia Ridgeway porphyry deposit and measure its persistence in the crushed rock produced by the Ridgeway block cave mine over one month. The model is then extended to demonstrate the effect of varying head grade, mixing and pod (or diversion) weight on bulk sorting efficiency.CITATION:Coghill, P J, Miljak, D, McGann, A, 2016. Modelling Grade Variation in Crushed Ore at Cadia Ridgeway and Its Potential Consequences for Bulk Ore Sorting, in Proceedings The Third AusIMM International Geometallurgy Conference (GeoMet) 2016, pp 77–88 (The Australasian Institute of Mining and Metallurgy: Melbourne)."

Jun 15, 2016

By AIME AIME

TWENTY-TWO sections and all four of the divisions sent delegates to the annual meeting. They became so interested in the wide ranging dis6ussion of old and yet ever-new problems of Institute affairs that three instead of two sessions scheduled were necessary. Monday morning the meeting assembled with President Bassett presiding in the absence of President-elect Tally, who had been called to Baltimore by the serious illness of a brother. Thursday morning Vice-president Turner presided until Mr. Tally appeared after which the latter guided the sessions. Since reports from all of the active sections save one had been mimeographed and sent in advance to the delegates, detailed reports from each were dispensed with. The discussions dealt largely with membership problems, both local and national, with financing, speakers and programs, visits from national officers, and the conduct of the annual meeting. Many interesting and constructive suggestions were made and will be reported back to the sections by the delegates.

Jan 1, 1931

The character of alluvial gold deposits presents severe difficulties, and sometimes limitations, on the extent to which metallurgical efficiency can be measured. Very little monitoring of metallurgical performance is practised. An indication of performance is often extracted from the comparison of production to the estimate. This is a dubious practice. The recovery efficiency of conventional practices, viz sluices, tables and jigs, falls off sharply below 0.2 mm. Field evaluation procedures tend to simulate these conventional practices. Thus the presence of fine gold may not be detected, particularly if obscured by clay. There is increasing evidence that fine gold could, in places, be present in sufficient concentrations to be economic. Most gold in paleoplacers is finer than 0.1 mm. There are established practices elsewhere in the mineral industry that may physically recover gold down to 5 microns in size and which could find application for fine alluvial gold.

Jan 1, 1984

By Stevak D, Plesingerova B, Ellias K, Tkacova K

This paper presents information on the processing of minerals by decrepitation that can be an integral part of their mineral dressing. In the case of carbonates, processing by decrepitation is a special operation that has its objective as concentration of the oxide. Selective fragmentation of magnesite brought about by decrepitation uses pseudomorphosis which was observed at thermal decomposition of carbonates. By decrepitation of magnesite high-grade MgO concentrates were obtained with MgO contents above 80 per cent. With the sideritic ores a very important phenomenon was observed consisting in the removal of quartz in the fragmented fine-grained product which favourably affect pelletisation of the fines. The selective concentration of the MgO product brought about by thermal decomposition of dolomite accompanied with decrepitation was negligible.

Jan 1, 1993

By P Rubie, J Russell

This paper looks at the past, present and not too distant future of grinding mill relining.Since the adoption of autogenous grinding (AG) mills and semi-autogenous grinding (SAG) mills as the primary ore reducing devices in the comminution circuit, plant shutdown frequency and duration has largely been driven by the liner maintenance requirements of these mills.The past 30 years have seen the introduction of many technologies to reduce relining times and frequencies. Advanced relining systems have been developed which include single and twin Mill Relining Machines, capable of placing large (up to 8000 kg) liners, high energy recoilless THUNDERBOLT Hammers, and high capacity Feed Chute Transporters. Despite these advances, today many of the hazards associated with mill relining remain and reline times still vary considerably between mine sites. A sophisticated discrete event simulation program, designed to analyse the mill relining process, has revealed the dominant influences on reline times of crew performance and of the need for safe access to larger segments of the mill.A new generation of advanced and semi-automated technologies is currently being tested that will address these remaining challenges. These technologies remove people from the hazardous environment inside the mill and will allow techniques used in modern advanced manufacturing to provide a near-future step change in mill relining safety and performance.CITATION:Rubie, P and Russell, J, 2018. 21st century relining – the path to greater mine site profitability, in Proceedings 14th AusIMM Mill Operators' Conference 2018, pp 621–640 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Aug 29, 2018

By C. S. ROBINSON, T. M. Romslo

This investigation was conducted jointly by the Bureau of Mines and the Geological Survey to obtain further information on the source and economic possibilities of the copper mineralization exposed on the Copper Giant group of claims, Ajo mining district, Pima County, Ariz. The Copper Giant group of claims was first brought to the attention of the Bureau of Mines by Joseph Hunter (deceased), Phoenix, Ariz., part owner of the claims. In December 1942 these claims, and an adjoining group, were investigated by T. C. Denton, mining engineer, formerly with the Bureau of Mines, and Nels P. Peterson, geologist, Federal Geological Survey. It was concluded the area could not be considered as a source of copper for immed- iate emergency needs and that further investigation of the area, at the time, was not warranted. In the spring of 1948 the lessee, J. H. Greenway Albert, requested Bureau of Mines assistance in developing the property. J. H. Hedges, former chief, Tucson Branch, Mining Division, and Lincoln A. Stewart, mining engineer, Bureau of Mines, made a cursory examination of the property and noted that at numerous places, low-grade copper mineralization appeared on fractures and bedding of the Locomotive fanglomerate and Ajo volcanics, which are thought to be of Tertiary age. During the period of May 11 through 14, 1949, Bureau of Mines engineers3 mapped and sampled two areas of mineral- ization. On June 2 and 3, 1949, the areas selected by the Bureau and three other nearby mineralized areas were investigated by C. A. Anderson and R. S. Cannon, Jr., of the Federal Geological Survey. As a result of these examina- tions, further investigation of certain areas was believed to be merited. The Copper Giant area was selected as the most favorable in which to carry out exploratory and development work. The field work on the project was started in October 1949 and completed in December 1949. It consisted of road building, trenching, sampling, topo- graphic and geologic mapping, diamond drilling, and logging of core. Samples taken on the project and during preliminary examinations totaled 64, of which 63 were surface samples. Diamond drilling, started October 10 and completed December 7, consisted of two holes for a total depth of 1,542 feet.

Mar 1, 1952

By Georgette Hlepas

How often phrases like these have been made by project managers and project teams trying to explain why there has been cost growth, project delays and differing site conditions claims. But does it really take x-ray vision? Do we really approach geotechnical construction projects unprepared for the subsurface that may be encountered because we can't get the answer? Too often our industry seems to think the answer is, "yes." However, this is not because we can't understand the subsurface. All too frequently it is because a choice is made to save money or time by foregoing or shortchanging upfront subsurface investigations. This is accepting a risk at the beginning of the process and realizing it during construction. Case histories have proven time and time again that upfront investment in site exploration and constructing a ground model mitigates risk in design and construction. But how much do we prepare? What level of investigation (or investment) is appropriate? The answer of course, is it depends; how much risk are you willing to take for your project and what are the potential consequences?

May 1, 2022

By R. Cammack

"A fundamental component of the Hoek–Brown empirical rock failure criterion is assessing an appropriate mi constant. In developing their empirical criterion Hoek and Brown normalised triaxial test results by dividing the principal stresses at failure by the uniaxial compressive strength to provide mi values for typical rock types. Experience shows that Hoek triaxial tests for individual lithology types commonly have a large inherent scatter. In addition, tests are often conducted on core from random locations, which although being of one parent rock type, display variations in density as well as minor variations in grain size and alteration.This paper presents experience from projects the authors have been involved in which large sample populations of triaxial and UCS tests on similar rock types were available. Geophysical logging data was also available for some projects. The methods used to estimate the compressive strength for each sample and the effects on normalising the data sets are described. The variability of the data and uncertainties and assumptions involved in the process of assigning a mi value from the normalised data are also discussed."

Jan 1, 2015

Jan 1, 1923

Jan 1, 1923

By Kevin J. Ashley

Coal is deposited at a mine site in several thin seams which outcrop in rolling hills and dip at about 8° to the northeast. The deposit is located in the northeast area of the State of New Mexico in the Southwest of the United States. A drilling program was undertaken and was quite irregular. Not all of the seams appear in all of the drill holes. Some of the holes did not go deep enough, while others were drilled below the outcrop of some seams. Occasionally a seam would pinch out and not be seen in a drill hole. Also, the outcrops of the seams are buried in many places by several feet of alluvium. All of these factors make correlation of the seams between drill holes difficult. The drawing of cross sections in three directions and at many locations across the mining area have been necessary to assure the accuracy of the input data.

Jan 1, 1982

By J. R. R. Bowling, A. McFadden, M. Francoeur

The abstract discusses the increasing use of extensive duct-fan systems in underground hard-rock mine development. It highlights the challenges of designing these systems, particularly the risk of partial vacuum and the need for careful booster fan positioning. The paper explores factors influencing the design of these systems, including component limitations, operational wear margins, and power constraints. It also defines the concept of partial vacuum and its impact on ventilation ducts. The authors propose an algorithmic approach for specifying the location and blade pitch of in-line booster fans to prevent system failure.

Mar 17, 2022

By P. J. Bartlett, H. A. D. Kirsten

A two-dimensional numerical scheme for the simulation of the plastic stress-strain properties of rock, and the various components of rock reinforcement and shotcrete, was used to rigorously determine the characteristics of tunnel support under squeezing conditions. D-shaped tunnels varying in size from 2,8 to 6,5 m in height and width were considered. The four main mining stages in the BAS block cave on Premier Mine were simulated as alternative sets of field stresses. The rock reinforcement comprised 1,8 m long fully grouted rockbolts and 6,0 m long fully grouted cable anchors at various spacings, and of various tensile and bond strengths. Various thicknesses of shotcrete and footwall support configurations were also considered. It was found that shotcrete and rock reinforcement, owing to inherent differences in stiffness, do not function in unison. As monolithic application, shotcrete has an overriding avidity for the imposed loading and destructs itself unless of extraordinary thickness. Under squeezing conditions, it accordingly supports the excavation surface while the rock reinforcement contains on-going convergence. The loads in the various support components can be determined only by simulation of the shotcrete as a stress-relieved medium to represent its broken condition in the field. The loads in the rock reinforcement vary around a tunnel and are the largest in the sidewalls. The loads increase with an increase in tunnel dimensions, the increase in load in the sidewall due to an increase in height being larger than that in the hanging- and footwalls due to an increase in width. Depending on the strength of the rock, the loads in the rock reinforcement become prohibitively large beyond a certain size of tunnel. The reinforcement of footwall rock can be omitted without adversely affecting that in other walls, provided the tensile and bond strengths of the reinforcement are not limited. The resulting footwall heave may, however, result in undermining of the reinforcement in the adjoining walls. A relative depth of anchor of 1,37 provides adequate bond strength in the larger sizes of tunnel, whereas, in smaller tunnels, anchor length is determined by the allowable convergence. The lack of fun column grouting adversely affects the efficiency of rock reinforcement to a major extent. The loads in rock reinforcement are very sensitive to the strength of the rock. Owing to a lack in refinement and calculational rigour, empirical classification systems are an inappropriate means for the design of tunnel support in squeezing conditions and, as a result, grossly unclear-estimate the support requirements.

Jan 1, 1992

By A. L. Engel

In recent investigations of chloride Volatilization of gold at the Reno Station, it was found that the addition of small amounts of copper and NaCl to the charge brought about improved volatilization of the gold. The investigations described in this report were concerned with the recovery, in an aqueous medium, of fumes from chloride volatilization of gold. It was attempted to discover a suitable aqueous medium that would permit absorption, solution, or precipitation of all the volatilized gold immediately upon contact with the medium. A simple laboratory apparatus was employed to produce fumes by volatilization. To this furnace was attached an absorption train consisting of a series of chambers containing the aqueous medium, through which the fumes were drawn by suction. A satisfactory synthetic ore that contained metallic, gold in a quartz-matrix was made and utilized, enough. gold being present in a' small amount of material to permit an accurate, metallurgical balance of the volatized values' to be obtained in all the experiments.

Jan 1, 1949