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By L Neems

In conjunction with Orica Explosives, Worsley Alumina have managed a program of continuous improvement to their cap rock blasting over the last two years. Blasting conditions are challenging because of the hard and shallow nature of the material. Blasting costs must be controlled and the environmental consideration of noise and dust is critical. Worsley have sought to determine the optimum balance between drilling and blasting costs and fragmentation. As the cap rock has few inherent weaknesses, loader productivity is strongly controlled by the degree of fragmentation from blasting. The optimum blasting solution is also constrained by maintaining airblast levels at the surrounding residences below 115 dBL . The major blast design parameters optimised during the program were stemming type, stemming length, powder factor and blasthole diameter. A series of carefully managed trial blasts were fired during 1997 and 1998 to quantify the effect of these parameters. Fragmentation was digitally measured using OricaÆs Powerseive G technique. Shovel productivity, airblast levels and blasting costs were also quantified. As a result of the project, Worsley now have an increased knowledge of the key parameters driving blast performance, and their consequent impact on downstream processes, the environment and total mining cost.

Jan 1, 1999

By C. H. Schack

Extractive metallurgy techniques currently employed to recover silver from primary ores and secondary scrap were reviewed and evaluated to identify and delineate metallurgical problems whose resolution would increase the supply of silver from domestic materials. The survey revealed that practically all primary silver is recovered by froth flotation concentration of various types of ores, followed by smelting and refining of the separate lead, copper, and zinc concentrates. Silver losses during the complex processing are, with rare exception, inconsequential. Furthermore, any silver production gain from base metal ores, which contribute about two-thirds of the total primary silver, is directly dependent no improvement in the market for the base metals. Statistical data on new and old silver scrap are wholly inadequate to make reliable conclusions about scrap processing problems and the possibility of increasing the recovery of silver from the many forms of secondary silver-bearing materials. The available information indicates (1) that efficient silver reclamation practices are employed by the larger users of silver, and (2) that recovery of additional secondary silver involves solution of problems of collection rather than of processing. Several possibilities exist for limited increases in silver production from primary ores by improvement in metallurgical practices. Specifically, research is needed to (1) solve the problem of recovering silver from the copper ores of the Copper Range district, Michigan; and (2) devise economically feasible methods for recovery of silver from presently marginal or submarginal ores char because of grade, location, and frequently complex nature cannot presently be exploited by conventional techniques.

Jan 1, 1965

Reconciliation involves the collection of tonnage, grade (quality) and contained metal (product) data from disparate and hopefully independent sources. Examples are exploration data, production sampling data from blastholes or drawpoints, and process plant data. These data may be compared by means of ratios (factors). The F1 factor usually relates short-term (ore control) model tonnages, grades and metal content to ore reserves depleted. The F2 factor usually relates received-at-mill tonnages, grades and metal content to delivered-to-mill = production tonnages, grade and metal content. The F3 factor is F1 × F2 and enables a comparison of a mineÆs ability to recover the tonnage, grade and metal content estimated in ore reserves. The F1 factor measures the accuracy of orebody knowledge in the ore reserves to the demarcation of ore and waste by ore control (short-term model). The F1 factor may be used to check and calibrate the selectivity of mineral resource models and/or planned dilution assumed in transfer from mineral resources to ore reserves. The F2 factor enables a check on unplanned dilution entering the ore stream between ore control and the mill. Using the factors it is possible to calculate a monetary value on improvements in the accuracy of orebody knowledge, selectivity and the effects of dilution and ore loss. Reconciliation should be an implicit part of the mining process, and reconciliation targets should be a key performance indicator for well-run mines.

Sep 26, 2011

By D. R. Willmer, E. J. Haas

"DISCLAIMER The findings and conclusions in this paper are those of the authors and do not necessarily represent the views of the National Institute for Occupational Safety and Health. Reference to specific brand names does not imply endorsement by the National Institute for Occupational Safety and Health. ABSTRACT This study analyzed the publicly available health and safety policies of 26 mining companies to determine the quantity of health and safety (H&S) practices that mining companies encourage in relation to the Plan-Do-Check-Act cycle. A thematic content analysis of the policies identified elements and practices within the text. On average, companies communicated information on about 7 elements (range 1–14, standard deviation = 3.49) and discussed 15 practices (range 2–34, standard deviation = 9.13). The elements in which companies highlighted the most practices were risk management, emergency management, leadership development, and occupational health. A discussion of the policy trends shows areas that mine sites can improve upon within their Plan-Do-Check-Act cycle, in addition to encouraging the use of both leading and lagging indicators to manage health and safety performance. INTRODUCTION Mining companies and their employees are expected to identify and manage risks at their worksites to ensure the health and safety (H&S) of everyone at the site. The primary mechanism in place to control these risks is a company’s health and safety management system (HSMS) (Boyle, 2012). Broadly, an HSMS is a set of standard, interrelated, and interacting elements used to promote and achieve occupational H&S goals (ANSI/AIHA Z-10, 2012; BS OHSAS 18001, 2007). The practices conveyed within an HSMS are expected to minimize incidents, injuries, illnesses, and even save worker lives (Alsop and LeCouteur, 1999; Arocena and Núñez, 2010). In the United States, both the Occupational Safety and Health Administration (OSHA) and Mine Safety and Health Administration (MSHA) regulate and encourage aspects of an HSMS (US Federal Register, 2011, 2010). Many HSMS documents (i.e., Chemical Industries Association, 1995; Health and Safety Executive [HSE], 1995; International Labour Office [ILO], 2001) attribute the source of their basic management system to Deming’s Plan-Do-Check-Act (PDCA) model of continuous quality improvement and organize their practices within this cycle (Johnson, 2002). The PDCA cycle is a well-adopted approach in health and safety management and promotes continuous learning and adaptability (Robson et al., 2007)."

Jan 1, 2018

By R. Vracar, D. Sinadinovic, Ž. Kamberovic

The paper presents a study on the oxidation of polymetallic gold ore by oxygen in an autoclave. It presents optimum parameters of autoclave leaching of sulphide refractory ore in self-generated sulphuric acid solutions aiming to get products with favourable technological characteristics from which precious and other useful metals can be recovered. A high partial oxygen pressure (1.5 MPa to 2 MPa), high leaching temperature (190°C to 200°C), pulp density 80 g/L to 240 g/L, ore fraction size 76 µm and leaching time up to 4 hours yield good results in rich ores. The full extent of copper and zinc conversion as well as a high extent of iron conversion has been achieved, while lead, gold and silver and a minor part of iron remain as a solid residue among gangue minerals.

Jan 1, 2001

Objective: Improve the dust control effectiveness of enclosed operator cabs on surface mining equipment.

Oct 1, 2007

By J. D. Christian

WHEN mining development began in the Kirkland Lake camp, twenty-five years ago, there was very little information obtainable from surface showings as to the nature of the ore deposits. A length of almost a mile of the main 'break' lay under the south arm of the lake, and elsewhere most of the outcrops were covered with glacial drift. Shafts were sunk on ore-bodies of uncertain value, either on the outcrop of the veins or-because of the lake to the north-in the south or hanging-wall. In the second year of the camp's development, the Great War broke out and retarded progress for its duration. He would have been a real optimist who, in those early years, could have foreseen mining at depths of more than a mile.

Jan 1, 1939

By M. D. Strachan

Technical & Procurement Services Limited (TECHPRO) in Ashford, Kent, UK, have been appointed as Project Managers for the major rehabilitation of the Copper Smelting complex of Nkana Division, Zambia Consolidated Copper Mines Limited, (ZCCM) Zambia. In outline, the overall scope covers three main areas: a) Transfer and implementation of new technology into Zambia for the production of primary copper. b) Replacement of gas collection and gas treatment circuits to guarantee high quality feed to the acid plants. c) Completion of a number of minor projects in the copper smelter and associated utility plants to complement changes in the primary smelting technique. This paper describes the Project Management methodology and style used by TECHPRO to tackle typical problems that arise in retrofit projects of this type. Reference is made specific aspects of the smelter projects to illustrate solutions to the problems. The subjects. addressed highlight the complexities not only of having a -Client 5,000 miles from the Project Office, but also the difficulty of co-ordinating and controlling the performance of min sub-contractors who are also geographically distant.

Jan 1, 1991

By D. Limpitlaw, A. Alsum, D. Neale

"Ecological footprint analysis (EFA) was first described in 1996 as a measure of carrying capacity appropriated by human activities. EFA is a resource and emissions accounting tool designed to track the demand on the biosphere’s regenerative capacity. Ecological footprints are increasingly used as indicators of organizational and corporate environmental performance and product sustainability. There is a compelling argument to develop an EFA tool for mining companies.To determine the size of a mine’s ecological footprint, land requirements for all categories of consumption and waste discharge must be summed. This land is only made up of the ecologically productive land and water in various classes (cropland, pasture, forests etc.) required on an ongoing basis to provide all energy and material resources consumed and absorb wastes.A challenge in conducting an EFA for a mine is the shortage of accessible data. Undertaking an EFA for a corporation or individual site entails compiling consumption and emissions data (which can be used for other reporting applications). The footprint results themselves highlight the most critical aspects of an organization’s impact on the environment and provide a platform for focusing actions and for educating the workforce to improve their contribution to best-practice operations. This paper discusses the benefits and challenges of undertaking an EFA for a mining company, and provides examples of the various components of ecological footprints associated with mines, as well as showing how an EFA can be used to understand and communicate some of the impacts associated with mining activities.IntroductionFundamental biophysical and economic limits are now being experienced in many countries around the world. Herman Daly, a former World Bank economist, described this scenario as ‘uneconomic growth’, where the costs of growth exceed the benefits (Daly, 2005). This occurs when the economy’s expansion encroaches excessively on the surrounding ecosystem, sacrificing natural capital. Under such conditions the sacrificed natural capital is more valuable that the resulting economic growth. Consequently, the world faces largescale threats to sustainability and especially to the viability and continued existence of the ecosystems that support human settlements (El Zein et al., 2014). In recognition of this sustainability imperative, mining companies are increasingly concerned with their ecological footprint (EF)."

Jan 1, 2017

By W. J. Sharwood

Cupellation is well known to be one of the most effective methods of separating silver and gold from base metals and other impurities, as well as one of the most accurate means for their estimation. In the latter application it consists in absorbing in some porous medium the fused lead oxide formed by the oxidation of the lead "bath," which carries with it the oxides of other base metals present. It has long been recognized that small amounts of both silver and gold were removed and absorbed together with the base metals, as well as being to some slight extent volatilized, and in work of the highest accuracy a correction is made for the "cupellation loss." Thus in the assay of bullion this is done by means of "proofs" of like composition, which are manipulated throughout in precisely the same manner as the assay pieces, and are assumed to lose or gain like amounts. In the case of certain very rich ores and commercial products the loss is adjusted by assaying the cupel itself and recovering the precious metal absorbed, but this doubles the labor and does not remedy the loss by volatilization. Arbitrary tables have been drawn up by various authorities with the object of enabling a correction to be applied without the use of a check. These have been of very limited practical use owing to the fact that a number of factors affect the losses sustained. For a fixed amount of precious metal the loss varies with the amount of lead used, with the nature and amount of the impurities, with the porosity of the cupel, with the air supply, and above all with the temperature at which cupellation is carried on. Pure silver loses relatively much more than gold, but the loss in either is diminished by an addition of the other metal. The concentration of precious metal in the litharge increases as the concentration in the lead increases, but little work has been published connccting the two. If all other conditions remain the same the actual total loss increases, but the percentage loss diminishes, with an increase in the weight of precious metal treated. The results of an immense number of experiments have been published at various times- during the past hundred years giving the actual or the percentage losses suffered under various conditions of cupellation, but the data are widely scattered and few attempts have been made to correlate them, so that no definite law has heretofore been shown to exist

Jan 1, 1916

By Jorge Cardenas, Diego Sanchez

Maximizing SAG mill throughput and metallurgical recovery in hard rock operations requires achieving target fragmentation, particularly a sufficient proportion of fines (<25.4 mm). These fines thresholds are strongly influenced by rock mass competency, commonly characterized using the Axb parameter (30–70) from the JK Drop Weight Test. In many operations, conventional blasting designs and standard emulsions fail to generate the fines content required for efficient downstream comminution. This study evaluates several advanced blasting strategies aimed at enhancing fines generation in very competent rock masses while maintaining cost-effective drill and blast performance. The first approach focuses on steadily increasing the powder factor (kg/t) as a core strategy, complemented by adjustments in blasthole diameter, reduced burden and spacing, the use of dual initiation points, short inter-hole delays (<25 ms), and optimized sequencing—all aimed at enhancing energy distribution and maximizing fines generation in highly competent rock masses. The second strategy consists of increasing the energy factor (kJ/t) by selecting explosives with higher energetic content and detonation performance. This includes the use of heavy ANFO blends and ultra high-energy emulsions with aluminum-based formulations, achieving Relative Bulk Strength (RBS) values between 180–190 and up to 210–230, respectively. These products offer higher detonation velocity (VOD) and peak pressure, enhancing shock energy transmission and breakage efficiency. Field results in granodioritic porphyries with high uniaxial compressive strength showed a 3–5% increase in fines generation compared to standard emulsions—contributing directly to meeting mill performance targets in high Axb conditions. Combined, these strategies provide a practical and technically validated pathway to achieving fines generation targets required for optimal mill performance. When matched to site-specific rock mass characteristics and integrated with blast monitoring tools, they enable improved fragmentation control, reduced P80, and enhanced plant throughput—without compromising operational efficiency or increasing costs disproportionately

Jan 26, 2026

By R Justa Cámara, L M. Pinillos Lorenzana

Barcelona Metro Line 9 (Line 9) is an essential infrastructure, improving the transversal transport in Barcelona by connecting the greater part of the currently existing radial lines. The purpose of Line 9 is to create a peripheral line surrounding the city of 1.6 million people (metropolitan area exceeds five million), and includes 52 stations, of which a total of 32 are deep shafts and 19 transfer stations in a densely populated area with several geographical constraints.The tunnel has an inner diameter of 10.90 m, bored with an earth pressure balance (EPB) tunnel boring machine (TBM) in soil sections with precast concrete lining type universal ring and segments 40 cm thick.A single tunnel solution was adopted, driven by functionality and costs savings. The internal tunnel section is divided into two levels to allow for one way rail transit including the platforms in each station on each level in order to optimise the space created by the TBM excavation and comply with the operational needs. The division in two areas of the tunnel was carried out through the construction of a horizontal slab resting on the tunnel segments with a constant thickness of 40 cm at the central part of the slab and supporting side walls.CITATION:Justa Cámara, R and Pinillos Lorenzana, L M, 2014. Two transit levels in one tunnel tube – an innovation in Barcelona Metro Line 9, in Proceedings 15th Australasian Tunnelling Conference 2014 , pp 387–398 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Sep 17, 2014

By F. Ebbutt

The copper-bearing waters now treated successfully at Britannia for the recovery of their copper content are almost entirely derived from what is known as the Fairview mine; this was the first extensive copper producer of the Britannia ore-bodies, and consisted of a number of almost parallel veins, carrying good copper values, the ore consisting of chalcopyrite and pyrite in a quartz gangue. These veins occur in what is known as the ?Britannia shear zone&apos; (Schofield, S. J., G.S.C. Sum. Rept., 1918, Pt. B), developed for roughly five miles in a roof-pendant in the well-known Coast Range batholith. This .pendant is made up of Palxozoic rocks known as the &apos;Britannia Formation&apos; (LeRoy, O.E., Portion of Main Coast of B.C. and Adjacent Islands; G.S.C. Memoir No. 996, 1908). These consist of dark carbonaceous slates, quartzites; and tuffs or basic flows, and within them are large quartz porphyry siJls. These Britannia rocks are sheared and altered, the resultant rock being a fissile quartz-sericite or quartz-chlorite schist. The mineralization invariably occurs in this chlorite schist, which in the more productive areas has mottles or blobs of chlorite developed. A narrow, though persistent, band of grey slate forms the footwall of the productive portions of the shear zone.

Jan 1, 1928

By Barnabas Bryan

In the year 1875, the Pennsylvania Supreme Court, acting on incorrect information concerning the production of oil and gas, established what has become the law of the land for the governing of oil production. While the decision was ideal for the refining monopoly which was completed in that year, there is no evidence that the court was in any way influenced by that refining monopoly. The evidence indicates that the court accepted the view of the royalty owner and independent oil producer. Throughout the life of the refining monopoly it accepted this law, which was so evidently to its advantage in securing cheap raw material. The same policy continued until the development of the exhaustion theory and the entry of the various units of the former monopoly into the production of crude oil. It is curious that no one doubted the new theory of shortage sufficiently to feel the need of changing the law. When the theory entered the economic field sufficiently to raise crude oil prices beyond justification it shortly disproved itself, but by this time the question of split ownership in producing fields made a change from legal error to justice too complicated to be undertaken by the large companies. Not even the ardent advocate of unit operation dared to carry the question to the United States Supreme Court, which has never been asked to pass on the complete question. The impossible position of the producing industry, hemmed in on one side by court law compelling wasteful production and on the other by the national law against cooperation, lasted until the demoralization of 1927 led to cooperation within a state and under state guidance. Since that time control of production has grown with the bringing in of each new field and at this time has the active support of the Department of the Interior in such ways as are definitely legal. As in 1875, the royalty owner and independent producer are opposing by all possible means the rational conduct of the business of producing crude oil. Fluctuation in Investments The investor in oil securities has not been as badly fooled as has the oil man himself. Fig. 1 is designed to show the composite regard of all

Jan 1, 1930

By Peter Chmara

The measurement of dust concentration in Ontario mines since 1926 has been accomplished by konimetry, based on South African experience and methods. The introduction of a voluntary program of semi-annual surveys of air conditions at all operations throughout Ontario mines created a need for an instrument that would assess konimeter spots rapidly and accurately. The development of the Electronic Scanner, it is felt, answers this need. A program whereby all konimeter spot samples taken at Ontario mines will be counted at the MAP AO field offices in North Bay is described.

Jan 1, 1969

By S. K. Kawatra

Outokumpu Electronics has designed a new on-line coal slurry ash analyzer, based on an instrument developed at Michigan Technological University. The primary goal was to build an accurate instrument at the lowest possible cost. The instrument determines ash content from measurements of x-ray backscatter, iron Ka fluorescence, and low-energy y-ray absorption. It is capable of measuring the ash content with a standard error of 1.1% ash over an analysis range of 9-21% ash and 7-22% solids. The analyzer was tested on coals from United States, Poland, Great Britain and the USSR.

Jan 1, 1992

By M. Paricheh, M. Rahmanpour

"The cost efficiency and high reliability of semi-mobile or combined in-pit crushing-conveying (IPCC) and truck systems, compared to conventional truck-shovel systems alone, make them more attractive for use in modern mining operations. Semi-mobile systems, which are the most common systems in open-pit mining, combine the advantages of both systems such that fewer trucks are required, operating costs are lowered, and the operation is environmentally sound. One major aspect of utilizing this system is determining the location of the in-pit crusher and the timing of relocations. Facility location models are rarely used in the mining sector, but these models may be used for solving the in-pit crusher location problem. To this end, the main parameters affecting the IPCC location in open pit mines are reviewed and examples given of how they are applied to a dynamic location-relocation problem. Subsequently, the model is implemented for Sungun copper mine in Iran and the number and the exact time of relocations of the IPCC units determined. IntroductionNowadays, compared to the last century, metal prices are more volatile, average and cut-off grades have decreased, and stripping ratios increased. From a mining engineering point of view, many ore deposits with simple access and low capital and operating costs have been exhausted. The remaining deposits are located at depth, with no easy access, which requires high initial capital costs (Osanloo, 2012). Mining at great depths and high tonnages is possible only through open pit mining. Today, more than 80% of open pit mines in the world use a shovel-truck system for loading-hauling operations (Osanloo, 2010).With increasing depth, haulage distances increase and the number of loads per truck decreases. As a result, fuel, tyre, and depreciation costs per ton increase. Owing to these factors, open pit mining at great depth (300 to 1000 m) using shovel-truck systems faces some technical and economic problems. In-pit crushing-conveying (IPCC) systems have been known in the mining industry for many decades. The idea was introduced in 1956 in Germany. Today, the cost efficiency and high reliability of IPCC make it more appealing than conventional shovel-truck operation, especially in longer life projects with lengthy transportation distances and high production rates (Koehler, 2003). Since 1956, different aspects of this system have been reviewed by many researchers. Barua and Lanergan (1985) developed a computer program that compares the tonnage aspects of various conveyor layouts. They believed that to maximize the economic benefit of lower haulage costs associated with in-pit conveying, one must first minimize the cost of the flatter slope, which is expressed in terms of additional stripping or tied-up ore. Dos Santos and Stanisic (1987) evaluated the design of a high-angle conveyor at the Majdanpek copper mine. Then, Sturgul (1987) applied GPSS (general purpose simulation system) to find the best location of an in-pit crusher."

Jan 1, 2017

By Norman Weiss

MOST papers on xanthate have dealt with principles rather than practice. On the assumption that many millmen are interested in knowing where and in what manner the xanthates are being used in mills other than their own, this paper will summarize the results of a recent survey of North American plants. A brief background story and a description of the xanthates will be followed with the results of that survey; then a few observations will be made on storage, feeding, and ore testing. The first xanthate was made in 1822 by Zeise of the University of Copenhagen. With commendable scientific detachment (or perhaps a bad cold) Zeise ignored a more pervading quality of the substance and named it for its color, after the Greek word "xanthos," meaning yellow. During the first 80 years of its recorded existence xanthate found use only in.small quantity as an insecticide; after the turn of the century there was some further demand for it in the vulcanization of rubber. It was not until 1922, by coincidence the centennial of Zeise&apos;s discovery, that C. H. Keller of San Francisco, seeking sulphidizing agents, recalled xanthate in connection with his experience years earlier when he had seen this material used in Europe to combat grape phylloxera. His successful tests with the new flotation reagent culminated in the basic xanthate patent (U. S. 1,554,216) in 1925. In the minds of many metallurgists the introduction of xanthate as a flotation reagent is identified with the beginnings of alkaline circuits, the first use of chemical collectors, and the beginning of selective flotation. Actually, these preceded xanthatc by a number of years. The alkaline circuit in lead-zinc flotation was introduced by Lyster as early as 1912. In 1922 Sheridann and Griswold made their momentous discovery (Timber Butte) that alkali cyanides and zinc sulphate effectively inhibited the flotation of pyrite and sphalerite in pulps made alkaline with sodium carbonate or bicarbonate. The use of lime to retard pyrite in copper flotation began within two years after Ralston directed attention to its possibilities in 1917. These were the beginnings of modern selective flotation, which was practiced successfully on copper-iron ores at Nacozari in 1922, and at Inspiration, Cananea, and others only shortly afterward. In the lead-zinc mills, selective flotation was of even greater economic importance at that time. The Sunnyside plant of U. S. Smelting, Refining, and Mining Co. used the process earlier than 1920, and Consolidated, Timber Butte, and others soon followed. As a chemical collector xanthate was relatively a latecomer, for the Corliss patent of 1917 covered the first of these, alpha-naphthylamine, and the Perkins patent of 1918 the second, thiocarbanilid. Both of these were in extensive use before Keller&apos;s xanthate made its commercial debut in 1923. There is general agreement, nonetheless, that the introduction of xanthate was

Jan 1, 1947

By S. W. Marcuson

Smelting of nickel-copper sulfide ores at Vale Inco’s Sudbury operations commenced in 1888 and the first converting plant went into operation in 1892. As the demand for nickel grew, the smelting operations expanded and by the 1960’s the Copper Cliff Smelter featured the longest converter aisle in the world with space for nineteen Peirce-Smith converters. During this long period, different converting techniques and vessels were employed. The evolution of converting is reviewed with special emphasis on the modifications made to the Peirce-Smith vessel to accommodate the metallurgical and process needs of the changing times.

Jan 1, 2009

J. E. Stead, Middlesborough, England (communication to the author): Prof. Howe&apos;s valuable paper on cast-iron brings forward most prominently the correct explanation of the part played by combined carbon in pearlite and cementite, in determining the strength and hardness of cast-iron. On a previous occasioil I have shown that castings made by melting a white Cleveland iron and glazed iron, one containing 1.5 and the other from 4 to 5 per cent. of silicon, and each about 3 per cent. of carbon, were stronger than those made of ordinary foundry-iron; the difference in the final castings being a differ-

Jan 1, 1902