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By A. W. Easton

INTRODUCTION This presentation will address changing trade patterns in lead and zinc from the mid 1980's to the early 1990's. Lead and zinc deposits are developed where they are found in many parts of the world. As many are not in major consuming areas, there is significant international trade in both lead and zinc concentrates and in refined metal. Forecasts of mine production have been analyzed, primary and secondary refining capacity and future consumption levels have been estimated, to arrive at the changes in trade patterns expected during the next half decade. This paper will first address concentrate and metal trade patterns for lead, then for zinc, and finally some remarks will be made on captive mine/smelter agreements, on regional trade agreements, and on barriers to trade. I would like to note that. there is only a small international trade in semi-fabricated products such as zinc die-cast alloy, zinc oxide, lead oxide or rolled lead and zinc. Most of the world's zinc and lead trade is in concentrates or refined metal. LEAD TRADE PATTERNS Lead Concentrates Six countries are exporters of lead concentrates. Three countries are in the Americas - Canada, Mexico and Peru, with two in Africa - Morocco and South Africa (including Namibia), and Australia. Due to the limited growth in lead consumption and consequent minimum investment in new production facilities expected, mine and metal production levels and future consumption levels are not forecast to change much between now and 1990. Accordingly, the tonnage of lead concentrates exported will remain essentially unchanged from these six countries. In total they have exports of 480,000 metric tonnes (metal content)* and by 1990 this figure is forecast to rise only slightly with the only increase taking place in Mexico. The concentrates are exported mainly to Western Europe & Japan. Apart from these six main lead concentrate exporting nations, a significant change is expected in the United States. In the mid * All concentrate trade figures are based on metal content

Jan 1, 1986

By Michael J. Humenick

A laboratory experimental research project was conducted to evaluate the use of chlorine for the oxidative destruction of residual ammonia which may remain in ground water after in-situ uranium solution mining operations. The work tested the idea of the injection of high strength calcium hypochlorite solution into the mining zone to convert ammonia to nitrogen gas as a final clean-up process for ammonia removal from the ground water system. The paper details ammonia removal efficiency as a function of chlorine dose, reactant and product material balances, and how the concept may be used as a final ground water restoration process.

Jan 1, 1982

By Richard Keith Corbin

INTRODUCTION The purpose of this paper is to assist explorationists and entrepreneurs in deciding whether to invest in gold mining concessions in lesser developed countries (LDCs) and, if so, in deciding which LDCs provide the best opportunity. The first decision is difficult in the face of common notions that foreigners are not welcome - only their money; that the best gold deposits were worked out long ago; that even if you survive the mal- aria, and in spite of 51% control by a national partner, you manage to profit from a new discovery, they will tax your operation to death, or worse yet, nationalize it. Close scrutiny of what happened to foreign investors in LDCs during the “Mexicanizationt” of the 1930's through the "Zambianization" of the 1970's shows that many of these concerns are well-founded historically. But the trend of knowledgeable writers over the past decade seems to be away from this gloomy foreboding, and towards a cautious optimism, based on recently perceptible metamorphoses in certain LDCs' conditions and attitudes. If the explorationist or entrepreneur decides to seize on this apparent opportunity, the second decision -- which of these certain LDCs to invest in -- is doubly difficult, because of the formidable tasks of coming up with a rating system to provide rational comparison, and then gathering enough current, reliable information to apply. To assist in these tasks, a suggested rating system and case studies, respectively, follow a brief review of recent writings on the subject

Jan 1, 1987

By W. J. Schlitt

Introduction A major uncertainty in assessing the potential for heap-and dump-leach projects is how to determine the extraction-rate curve for the recovery of the mineral values from coarse ore. Such material could either be run-of-mine (ROM) or primary crushed ore. The problem with field testing coarse ore, especially for new projects, is the large scale and extended leach times needed to accurately determine the final extraction-rate curve. At least 5 x 103 to 5 x 104 t of representative ROM ore are typically required for a copper test heap, and much more is often used. Kennecott, for example, recently constructed a 0.9 Mt (1 million st) ROM test heap at the Bingham Canyon Mine in Utah. In such coarse ore operations, the ultimate level of extraction will require a leach cycle that can extend from several months to a few years. Quite often, project development schedules do not provide the luxury of mining such large quantities of material or running such long tests. Instead, test data are usually limited to results from column leach studies on relatively fine ore, often with a top size that does not exceed 25 mm. Maximum leach times are also short, typically less than a year before an initial decision is needed on project viability. Proposed method One approach to estimating the recovery from a coarse ore leach is to assume that the leach solution will have some ultimate penetration distance into the rock. Then, the final level of mineral extraction in this "leached rim" will equal the ultimate level of extraction identified in various testing programs. Obviously, if the radius of a given rock fragment is less than the penetration distance, that fragment will be fully leached at the end of the operation. With larger rock sizes, the percent recovery will fall off as the size increases and the fraction of unpenetrated rock mass increases. Such an approach sounds simple but is likely to be complex when applied to a real project. For example, the penetration distance will be a function of both the rock characteristics and the effective length of the leach cycle. The important rock characteristics include rock porosity, the degree of internal fracturing and the mode of mineral occurrence. With regard to the latter, penetration is likely to be greater if the leachable mineralization occurs on fracture surfaces or in veinlets, as opposed to fine grains uniformly disseminated throughout the rock mass. An estimate of penetration distance may be derived from column or heap tests by noting the depth of solution penetration into the larger rock fragments after three, six and 12 months of leaching. While the penetration rate is ore specific, something on the order of 10 to 20 mm/y may be appropriate for competent, primary copper (chalcopyrite) ore. For gold in tight quartz, the rate may be about the same. Copper oxide ores and gold that is hosted in a more porous rock matrix are likely to have penetration rates that are at least two to three times higher, and an even higher rate should be appropriate for uranium hosted in sandstone. As noted above, the length of the effective leach cycle is likely to be measured in years. On this basis, the ultimate penetration distance (dp) would vary from less than 50 to several 100 mm when a particular ore is leached to exhaustion. Several sets of mathematical manipulations are necessary to convert a rock size distribution and corresponding value of dp into an estimated extraction-rate curve. The first step is to break the ROM size distribution down into intervals and then calculate the radius for the mean rock size in each interval. This is shown in Table 1 for rock sizes up to 1.75 m (about 6 ft) in diameter. The next step is to calculate the volume of unleached core and the fraction of rock that is leached. This is done for the following three values of dp: 25, 100 and 250 mm. Results are shown in Table 2. The third step is to select the ultimate level of recovery that will be achieved in the fraction of material that is effectively leached, i.e., the outer zone that is penetrated by the leach solution. This is clearly a site-specific factor that can only come from metallurgical test results on representative ore

Jan 1, 1998

By -Ing. Alexander Dohmen, -Ing. Gunter Hagen

In order to improve the thermal environment in deep coal mines, two principles can be applied: firstly, the reduction of heat flow into the air stream, and secondly, cooling of the air current. First of all one should try to minimize the heat-flow rate into the air stream underground by the mining method itself. German coal mines prefer the caving system of advance longwall mining simultaneously on several levels with ascensional ventilation. The effects of descensional ventilation and different face ventilation layouts on the thermal environment are discussed, besides the effects of pneumatic and hydraulic stowing. Application of all mentioned possibilities together instead of today's common mining method would reduce the Basis Effective Temperature considerably. In the future, hydraulic haulage of the raw coal underground substituting belt conveying may be a mean to improve the thermal environment further on. If the working depth of a coal mine exceeds about 1 000 m, cooling of the air stream under-ground will be necessary in addition to the above- mentioned measures. Generally, in German coal mines different types of indirect-contact air coolers located nearby the face are used. Besides such so-called cooling coils, large-capacity spray coolers are used because they work more efficiently if a great cooling duty is demanded. Common designs of both air cooler types and their possible utilizations are demonstrated. Deep coal mines in the Federal Republic of Germany need up to 30 MW of total cooling capacity. Mostly, the needed chilled water is cooled by large-capacity refrigeration plants placed on surface, underground or in combination of the above locations nearby the shaft. Due to the increasing underground cooling demand, the distribution of the chilled water and the control of the air coolers become more important. The conception of a control system is discussed, which is based on the control of the single air coolers. All air coolers of a mine are controlled so as to optimize the supply water-flow rate with regard to the demanded cooling duty of the coolers. On the next upper level of the control system the whole chilled water network and finally the refrigeration plant is regulated. Present research work deals with the use of ice for cooling mines. There are some unsolved problems in this new cooling concept. Some aspects of making, transporting and storing ice are described. The use of ice may probably reduce the cooling costs of German collieries, which are at present approximately 4 DM per ton of saleable coal.

Jan 1, 1989

By B. M. Davis

The problem of assessing uncertainty in grade estimates has been a topic of research and debate since kriging was first introduced. Use of the kriging variance was originally touted as all that was necessary in order to produce confidence limits for grade estimates. Since then the prudence of using the kriging variance for confidence interval estimation has been questioned, and techniques such as the jackknife and conditional simulation have been proposed to develop "true" interval estimates. This paper compares the differences between interval estimates produced by the three methods and makes conclusions about the performance of confidence intervals produced by the different methods.

Jan 1, 1997

By Zhou Aimin, Xie Benxian, Song Chongkai, He Zhexiang, Zhang Changqing

Concrete pump is commonly used to transport paste and high-density backfill over relatively long horizontal distances. However, since the investment in concrete-pump and its operating cost are quite high, the application of paste backfill is limited in China. For seeking an alternative paste backfill technology, squeezed-transport technique and equipment has been developed. The new technology has been adopted at Tonglushan Copper Mine of Daye Nonferrous Metals Corp for overcoming a difficulty of long-distance transportation of high-density cemented tailings fill. It proved to be an alternative paste backfill method with simple backfill system, saving investment and low operating costs.

Jan 1, 2001

By Marco Sonnenschein

Belt conveyors for tunnelling are industry standard and key to high performance of TBM drives. Conveyors reduce traffic in the tunnel and increase safety in the tunnelling process. This paper will illustrate the state of the art of tunnel belt applications in very long and curvy tunnels like the Coxwell Bypass Tunnel Project in Toronto, Canada.

Jun 23, 2024

By Mustafa Kumral, Samer Hmoud

Effective management of ore loss and dilution is essential for successful grade control and short-term mine planning due to their significant impact on the economic, environmental and technical aspects of openpit mining operations. Factors influencing ore loss and dilution fall into two categories: (1) controllable factors, like mine equipment selectivity and blast design, and (2) uncontrollable factors, such as spatial heterogeneity of ore and waste blocks on a bench. This paper focuses on the second category by applying spatial entropy concept to capture heterogeneity at the scale of selective mining units. In this paper, global spatial entropy is used to assess the impact of spatial heterogeneity between ore and waste blocks on the magnitude of ore loss and dilution, while the local spatial entropy can guide the allocation of blast movement monitoring balls preblast. High values of the global spatial entropy indicate increasing potential of ore loss and dilution, which reduce profit. Furthermore, through a number of case studies, the study investigates the relationship between spatial entropy, cutoff grades, blast movement, dig-limits optimization model running time, and profit. The results show that changes in cutoff grade and blast movement can significantly affect spatial entropy post-blasting and increase ore loss, dilution and profit reduction, revealing the need for controlled blasting at specific bench sections. Additionally, the results demonstrate an exponential increase in profit reduction due to ore loss and dilution with a rising global spatial entropy

Jun 1, 2024

By R. W. Macqueen

Organic matter commonly associated with Mississippi Valley-type (MVT) lead-zinc deposits may be used to assess the burial/thermal history of host rocks and sulfides. Organic matter may also have been involved in the genesis of H2S required to precipitate sulfide ores. Pine Point, Northwest Territories, Canada, provides an example. Located on the south shore of Great Slave Lake, Pine Point is an operating world class deposit which has produced 58 million tons of ore containing 3.0 percent Pb and 6.1 percent Zn from more than 40 Individual orebodies. Well-documented paleokarst features in a Middle Devonian carbonate complex are the main ore hosts. Sphalerite and galena ores are mainly open space fillings, and, from fluid inclusion studies, were deposited from saline fluids at temperatures of up to at least 100°C. Forms of organic matter occurring in the Pine Point barrier-reef complex Include indigenous organic matter, heavy oil and bitumen. Based on numerous geochemical criteria, it is clear that heavy oils and bitumens were derived locally from indigenous organic matter. The geochemical character of the Indigenous organic matter demonstrates that the barrier-reef complex is immature to marginally mature with respect to petroleum generation, and has not been heated above about 60°C maximum. Thus sphalerite with fluid inclusions that yield temperatures up to about 100°C represents former thermal anomalies. Bitumen is of two main forms at Pine Point: widespread soft, tarry material soluble in organic solvents (unaltered bitumen); and hard, brittle, glassy material that is mostly solvent-insoluble, and associated with white sparry dolomite and lead-zinc ores (altered bitumen). There are also partly altered bitumens. Unaltered bitumens have atomic H/C ratios of about 1.4, sulfur contents averaging 7.8 percent, and sulfur isotope ratios 634S) of approximately +4.6 per mil. Altered bitumens have atomic H/C ratios averaging 1.02, sulfur contents averaging 22 percent, and 634S values of about +12.9 per mil. These data and the organic geochemical character of a large suite of bitumen samples analyzed, strongly support the concept that some bitumen has been involved in the thermochemical reduction of sulfate to produce the H2S required to precipitate the ores. Mass-balance calculations indicate that the amount and degree of alteration of bitumen is more than enough to account for the reduced sulfur species deposited as metallic sulfides at Pine Point. MVT deposits, including Pine Point, seem to require the coincidence of available metallic Ions, pre-existing pore space, and a precipitating agent. A close look at organic matter in more MVT settings may outline trends in distribution or alteration which can be used as indicators of past thermal environments and processes of ore deposition.

Jan 1, 2013

By Mokhinabonu Mardonova, Yosoon Choi

Open-source technology to support Industry 4.0 has become the target of academic research in many sectors. This study reviews the trends of open-source technology, its forms, and some industrial applications. As a case study of the use of open-source hardware and software in the mining industry, this study proposed a low-cost environmental monitoring system for nonmetallic underground mines to support mine safety and occupational health. The system was developed using open-source hardware, Arduino and 3D printer, to design the monitoring device, and open-source software, MIT App Inventor, to develop an Android application for smartphones to enable remote communication with the system. Field experiments were conducted at an underground tunnel and a nonmetallic underground mining site to assess the performance of the system in mobile and static modes, respectively. Although a few limitations related to the precision of the low-cost dust and gas sensors exist, the findings of the experiments show that the mining industry can benefit from open-source technology deployment when considering cost factors.

By S.K. SHARMA, ANKUSH GALAV, G.S.P. SINGH

This paper proposes a hydromechanically coupled numerical modeling approach to quantify the influence of various parameters on the mechanical and hydraulic performance of protective water barrier pillars (PWBP). The results are validated for two case studies having pillar widths of 16 and 42 m at cover depths of 136 and 190 m and water heads of 25 and 141 m. The outcome can be significantly helpful in the rational design of these pillars in underground coal mines.

Mar 1, 2023

By MOHAMED WAEL, ANASTASIA GOEVA, ABDULAZIZ TAWFEEQI, IHTISHAM TARIQ, S. M. ZAKIR HOSSAIN, NIAZ ALI KHAN, MUHAMMAD FAISAL IRFAN

Carbon dioxide (CO2) mineral carbonation is a carbon capture and storage (CCS) technique that can sequester CO2 permanently in the form of durable carbonate minerals. Extendable permanent storage capacity — onshore and offshore storage reservoirs — and the exothermic nature of the mineralization process make this technique attractive [1]. In the present study, maximum carbonation efficiency for aqueous mineral carbonation was achieved through two steps: optimization of calcium leaching from cement kiln dust followed by the reaction of pure CO2 with the calcium hydroxide precipitates formed by hydroxylation using sodium hydroxide (NaOH). Optimization of calcium leaching is highly important, as it is a rate-limiting reaction step in mineral carbonation and also influences and enhances the carbonation efficiency. To maximize the calcium leaching as well as carbonation efficiency, response surface methodology — with either Box-Behnken design or central composite design — was applied. It is the most well-known statistical, fully randomized and bias-free method for determining optimized conditions using quadratic regression models.

By Wedam Nyaaba, Edward Wellman, Angelina Anani, Eduardo Cordova, Sefiu O, Cluber Rojas, Adewuyi

This study aims to investigate the optimal method for ore pass construction in incompetent rock masses. We evaluated the conventional and raise boring methods based on safety, efficiency, excavation control and ground support for ore pass construction. A stability analysis using analytical Q-raise (QR method) and kinematic analysis methods for ore pass construction with a raise borer before and after grout injection of the rock mass was performed. As a case study, an ore pass (diameter of 3 m, depth of 100 m) within an incompetent rock mass was considered to gain further insight. The rock mass was characterized according to the classification methods Q Barton, rock quality designation (RQD), rock mass rating (RMR) and geological strength index (GSI). The grout intensity number (GIN) method of grout injection was used. The safety factor of less than 1.075 obtained before injection was lower than the acceptance criteria. However, grout injection before raise borer excavation resulted in a safety factor greater than 1.5. Using a raise borer without pre-grouting indicated that the maximum unsupported diameter (MUSD) of the ore pass was less than the required 3 m and equal to or larger than 3 m post-grouting.

Dec 1, 2024

By Ralph M. Barnett

It is no longer a question of "If" mining operations will meet standards and regulations set by various governing bodies and the demands of citizen groups. Legislation is fact. The needs of citizens are becoming more of a integral part of mining company planning. "Open Planning", regardless of some frustrations it provides, is working. Climax, AMAX Chemical and AMAX Exploration are successfully employing several citizen involvement techniques in exploration and feasibility study phases of planning for potential new mines.

Jan 1, 1978

By Adrian E. Scheidegger

Recently, a new theory of the subsidence of rock masses has been advanced by Litwiniszyn and coworkers, which is based on the notion that the individual rock particles perform random walks. It is shown in the present paper that the final equations of the random walk theory can be justified from general principles of statistical mechanics in which no specific theoretical model of particle behavior has to be made. The theory appears to account appropriately for subsidence phenomena.

Jan 1, 1966

By Gilman Y. Murray

The problems connected with the development of economically acceptable water supplies in arid regions are discussed. The techniques for capture and retention of runoff from infrequent desert storms are described and the major factors involved are emphasized. The exploration and planning for a typical water development in southwest United States is described, including an analysis of capital and operating costs for the proposed system.

Jan 1, 1963

By Silas Li, Michael Harris-Gifford, Jason Choi, William H. Hansmire

Twin-bore tunnels with a single track in each tunnel are currently the most common subway tunnel configuration. However, a large diameter single-bore tunnel with two tracks separated by a center wall could replace twin-bore tunnels and introduce potential benefits but require necessary modifications to the subway system. In particular, a nominal 40-foot-diameter single-bore tunnel can be considered in lieu of nominal 20-foot-diameter twin-bore tunnels. This paper discusses the benefits and modifications introduced by replacing twin-bore tunnels with a single-bore tunnel, with respect to various elements of the subway related to the tunnel, stations, track, and tunnel systems.

Dec 1, 2022

By D. R. Zeni

Current mine development requires the safe, 'expedient, economical construction of vertical shafts for ventilation and access in diverse ground conditions. Advancements in large diameter shaft drilling have lead to exceptional mechanical refinements of shaft drilling technology. The author traces the history of the equipment and methods employed by the large diameter shaft drilling industry and explores likely paths of future development. A few case histories are cited to illustrate the adaptability of shaft drilling as a versatile tool for advanced worldwide mining operations. Safety, schedules, cost, environmental impact, and capability are discussed to assist the mining engineer in the investigation of feasible projects.

Jan 1, 1994

By M. B. Mooiman

Cyanidation of gold and silver ores by conventional mining and milling techniques has been practiced for over a century. During the past decade, heap leaching has emerged as a viable process for treating certain gold and silver ores. (1) The rapid development of heap leaching for gold and silver ores is largely a result of lower capital and operating costs and faster start-up time than that offered by conventional mining and milling. In general, these features make heap leaching an ideal approach for exploiting either small or large low-grade disseminated deposits of gold and silver considered to be uneconomic by conventional methods. The development of heap leaching technology has evolved from the operations initiated in the western United States. The basic components of a typical heap leaching operation are depicted in Figure 1. Run-of-mine or crushed ore (usually 1 to 5 ppm gold) is delivered to specifically prepared impervious pads. Alkaline cyanide solution is applied to the surface of the heap and allowed to percolate through the heap (typically 10-30 ft high). Pregnant leach solution which drains from the heap is collected and stored in tanks of sufficient size or in collection reservoirs. Gold and silver are recovered from the pregnant solution and the barren solution is recirculated back to the heaps after reagent make-up.

Jan 1, 1983