Search Documents

By J. J. Gusek

An 81 hectare (200 acre) tailings impoundment at a taconite operation in Wyoming that ceased operations in 1983 has been a source of blowing dust. The site qualified for reclamation under Wyoming's Abandoned Mine Land (AML) reclamation program. The ultimate reclamation construction included: incorporating commercially available organic amendments and fertilizers into a 300 mm (12 in) thick "cap" of a sterile clay loam cover material, lowering the water level in an adjacent flooded mine pit that was feeding uncontrolled seeps into the tailings basin, constructing a wide tailings pond spillway that allows flood control while minimizing seasonal water level fluctuations in the pond, and building rock berms and rock snow fences for wind erosion control and soil moisture preservation. Tree planting in the protective wind/snow shadows of rock berms and rock snow fences, and the plugging of a decant pipe are planned for future work.

Jan 1, 1998

By David Dreisinger

The cyanidation of copper containing gold ores results in the co-dissolution of copper along with gold. Copper dissolution results in elevated cyanide consumption that may make the treatment of the gold ore uneconomic. Copper-cyanide species in gold mill effluents may increase the environmental toxicity of the gold mill solutions. The objective of this study was to examine the use of solvent extraction to recover the copper cyanide species from gold mill effluents. Solvent extraction with Cognis (formerly Henkel) reagents LIX 79 and LIX 7950 was studied. Excellent extraction of copper and cyanide was achieved using pH control in extraction. Copper and cyanide may be stripped from the reagents using basic eluants. Various routes for copper and cyanide recovery from the strip solution were studied. Excellent results were obtained using electrowinning in both divided (Nation membrane) and undivided cells. Cyanide recovery was obtained by AVR treatment of a bleed stream from electrowinning. Full results of solvent extraction, copper and cyanide recovery are reported.

Jan 1, 1998

By Tan Pengfu

By using computer-aided techniques of equilibrium calculations for multicomponent and multiphase systems, a computer model has been developed to simulate QSL, KIVCET and Outokumpu flash lead smelting processes. A mathematical description is studied using two zones or three zones, namely smelting zone and slag reduction zone for QSL furnace, or offtake shaft zone, smelting shaft zone and electric furnace zone for KIVCET furnace. The amount of each component in lead, slag and gaseous phases under equilibrium is calculated by using the data of free energies of formation and activity coefficients. The predictions by the present computer model are compared with the known data from QSL-Plant at Korea Zinc Co. in Korea, KIVCET-Plant at Nuova Samim Co. in Italy and Outokumpu Co. in Finland. The agreements between the computer predictions and the known data are excellent, so that the present computer model can be used to monitor and optimize the actual industrial operations of direct lead smelting.

Jan 1, 1998

Safety bulletins and stickers issued for the paste backfill plant at Cannington read as follows: æCannington Paste Fill Plant, First of ItÆs Kind in AustraliaÆ. We could even dare to state that the Cannington plant is different than any other paste backfill plant in the world because of the overall design, plant equipment and underground delivery. However, one thing to be certain of is, the paste backfill plant works as designed and a pump is not required to deliver paste underground given the general plant arrangement at Cannington. The past two years have been a process of evolution to design and construct the plant and make it work. This paper will describe the paste process using full mill tails along with placement of cemented paste underground. Discussions will include rationale of fixed plant selection along with various control requirements. BHP World Minerals recently commissioned its paste backfill plant in December 1997. Cannington is placing full mill tails underground via two surface boreholes that deliver either to primary or secondary stopes.

Jan 1, 1998

By Stephen Tadolini

The annulus thickness, for bolts anchored with resin, effects the systems final performance. Recent developments in the manufacturing process of these support systems, coupled with improved resin characteristics, have resulted in design changes to minimize hole drilling requirements and reduce resin usage. The amount of resin to properly anchor a 7/8-SRD (Special Rolled Dimensions) bolt in a 25.4 mm (1-in) diameter hole has been reduced to about one-third of the resin required to anchor a No. 7 rebar bolt (22 mm (7/8-in) diameter) traditionally installed in a 35 mm (1-3/8-in) diameter borehole. More importantly, laboratory and field tests, indicate that higher bond strengths and anchorage capacities can be realized with reduced annulus systems. These factors have a direct bearing on reinforcement efficiency and subsequent roof support performance.

Jan 1, 1998

By J. X. Guo, A. Pindred, J. Moore

"Two modified centrifugal flotation cells (CFC), Q2 and Q3 models, have been developed for froth flotation. The operating mechanism of Q2 uses a rotating bowl to create a centrifugal force on the feeding slurry. Froth concentrate is collected on the central top of the cell and tailing is discharged along the top edge of the rotating bowl. The Q3 utilizes a vertical rotating feed pipe to introduce the feed slurry into a stationary flotation column cell. A centrifugal force is also created by a rotating pipe for the feeding slurry. Froth concentrate floats over the cell and tailing discharges through the bottom of the cell. The centrifugal force causes the flotation recovery of fine particles from CFC to be higher than from the conventional cell. The major operating variables of CFC are the rotating speed, air flow rate into feed slurry, and slurry feeding rate etc. Flotation tests have been carried out on synthetic mixtures of galena-silica and chalcopyrite-silica, as well as some natural base metal ore samples. Parallel tests were also conducted on a conventional flotation cell. The comparative test results demonstrated that both Q2 and Q3 have the advantage of higher recovery of finer than 10 µm mineral particles, which are very difficult for the conventional non-centrifugal flotation cell to recover. The grade-recovery curves produced from CFC are equivalent to or better than the conventional flotation cell. There is a potential in the future for the CFC to be employed in slime flotation.1. IntroductionA centrifugal flotation cell (CFC) was originally developed by Clean Earth Technologies [I] to separate oil from water and was subsequently applied in froth flotation. Figure 1 illustrates the design of the original CFC, in which a vertical feeding pipe is attached to the center part of a rotating drum and a stationary shell encloses it to direct the float froth and tailing discharge. The slurry flows down to the bottom of the spinning drum, gets in touch with a screen of air bubbles and then migrates outwards under the effect of centrifugal force. After the mineralized froth enters the rotating cell, it is pushed back to the inner surface. The overflowing froth is collected in the froth launder while the tailing slurry is discharged through orifices along the outer rim of the drum."

Jan 1, 1998

By Julian Malnic

Transplanting the steps used in exploring and developing mineral deposits on land into the marine environment will help developers. While familiar to oil and gas explorers and many other sea-based industries, the marine environment is so alien to the 90s terrestrial miner that the phrase ?terrestrial mines in the sea? is useful for reinforcing the fundamental parallels in mining sulphides at sea and on land. As on land, the processes of ?pre-feasibility study? and ?feasibility study? will determine the economics of mining a deposit. For many, the step into the sea can detract from the need to follow these familiar and recognised steps used when any new deposit is found on land. The technology to mine sulphides in the 1500 to 2000 m depth range exists. Currently, it is not found in a single, integrated and commercially available system but every element available to build such as system is. Transplanting the thinking of the leaders of the terrestrial mining industry into the sea will perhaps be the greatest challenge and reinforcing the basic steps of discovery - pre-feasibility through to feasibility is the best way to achieve this. After all on land one doesn?t ask ?Can we profitably mine it?? before one can authoritatively answer ?What have we found??. Research by Nautilus Minerals Corporation suggests that the mining of seafloor massive sulphides (SMS) will not only be profitable but potentially lower in both capital and operating costs than prevail in most terrestrial deposits. Nautilus is also acutely aware that with the first SMS deposits under title, it will be setting the standards for seafloor mining into the new millenium. Indeed, we have trained on us not only the eyes of the mining industry waiting to see if SMS deposits really do offer radically better returns, but other stakeholders of the sea and the capital markets. Mistakes by any SMS developer could land all developers in ?hot water?. The rush to exploit our deposits for economic gain will be tempered by ensuring there is adequate public debate and consultation for our environmental code. Failure in any facet of a marine mining development could set back the move to the ocean floor.

Jan 1, 1998

By Daniel Hodouin, Mélanie Proulx, Claude Bazin

"Collector and frother are commonly used to induce the floatability of valuable minerals in flotation plants. These reagents are usually distributed down a flotation bank or sometimes added in a single location at the top of the bank. A review of the literature and some experimental results obtained from laboratory and full scale trials indicate that the way flotation reagents are distributed down a flotation bank may influence the flotation response of coarse particles. The paper summarizes these observations which show that the distribution of reagents is a potential tuning variable for process optimization.IntroductionFlotation is in many cases the only way to concentrate sulphide and some oxide minerals. Flotation is a concentration process that uses particle surface composition to induce the selectivity by mean of hydrophobic reagents. The influence of particle size on flotation was firstly analyzed in the 50's1·2 and intensively investigated in the 70's, early 80's and 90's3-18. Most of the observed results show the existence of an optimum flotation size range9 whose recovery is not significantly influenced by a reduction of collector dosage""18. This result brought the idea of using a small portion of the total collector dosage at the top of a rougher bank and to distribute the remaining through the end of the bank to ensure sufficient collection of the coarse particles that need more hydrophobic coverage to attach to air bubbles and be lifted to the surface.This paper discusses the idea of using the distribution of reagents down a flotation bank to tune the operation of the process. Unlike the literature on the effect of reagents (collectors, frother...) dosage that is dense, little attention has been put on the distribution of reagents in a flotation bank. Although research on this subject has not been pushed very far, some operators have recognized the potential of a strategic addition of reagents down a flotation bank19. Except for a few exceptions, the proportions of the reagents to be distributed down a flotation bank are decided according to operator's experience instead of process analysis in relationships with particle size. This paper recalls observations made by different researchers that lead to the idea of using the distribution of reagents as a tuning variable for the flotation process. The paper does not present a detailed analysis or optimization scheme for tuning the distribution of reagents, but analyzes some plant and laboratory results with a view toward the development of a testing methodology and ultimately a validation or rejection of the idea."

Jan 1, 1998

By Don McCarthy

Modern mining has been dominated by the drill-and-blast method for more than a century. There have been steady improvements in drilling technology, explosive products, methods of initiation and integration of the two techniques but the essential cyclical nature of drill-and-blast rock excavation naturally limits continuous improvements in cost and productivity. Dynamite and safety fuses, compressed-airdriven tools, and most of the crushing and grinding machinery in use today were all developed before 1900. Moreover, virtually every mechanized mining system used today was in use by the late 1950’s.

Jan 1, 1998

By William Bookshar

The development of increasingly more sophisticated and powerful retreat longwall equipment has led to ever wider and longer panels. To maximize gateroad development efficiency, provide for effective face recovery, andlor reduce ventilation pressures, entries are cut across longwall panels in advance of longwall retreat mining. These entries must be supported to prevent falls of immediate roof in advance of the face. Recently, cut through entries and predriven recovery room entries have been supported with various combinations of roof bolts and passive support. The cost and relative success of these support systems has varied considerably. Eighty-Four Mining Company, Fosroc, and Sub-Technical have developed and implemented an efficient, yielding, multiple component, cement crib to support longwall panel cut through entries and pre-driven recovery rooms. The cutable crib consists of an inexpensive flyash cement pillar poured in a bag, within a temporary metal form. The form is moved to a subsequent cutable crib location after the initial hardening of the flyash pillar. The remaining entry height between the mine roof and the top of the flyash cement pillar is filled with a specially designed bag filled with Fosroc's Tekset cement formulation. The diameter of the pillar and thickness of the Tekset bag are designed to accommodate the anticipated front abutment loading and entry convergence.

Jan 1, 1998

By J J. Laas, A J. S Spearing

Considering the capital outlay, running costs and inherent difficulties of backfilling operations at depth the benefits need to be clearly established, quantified and then exploited. Expansion of Vaal Reefs Exploration and Mining Company towards even deeper, highly faulted ground, and the process of establishing codes of practice for the management of rockfalls and rockbursts, required that a sound support strategy be formulated. This paper describes the derivation of an overall backfilling strategy on Vaal Reefs. Use was made of experience gained over some 12 years with backfilling. Numerical modelling was used to estimate potential benefits of backfilling the deepest mining areas still to be extracted. In exploiting a probably unique opportunity, two contiguous mining areas bounded by the same major geological structures, were compared. One was extensively backfilled and the other not at all. Shaft sinking through a backfilled stope added another dimension to the database.

Jan 1, 1998

By Luis Valentim, Terry Cousins

The goal of a mine is to produce ore at the lowest possible cost with due regard to safety of pemonnel and equipment. The mining process consists of a number of discrete operations such as planning, drilling, blasting,excavation, transportation and crushing. These processes are interrelated and cannot be considered as separate activities when optimising output and minimising cost. In order to optimise the global costs it may therefore be necessary to increase cost in one area such as blasting design and management beyond the norm.

Jan 1, 1998

By I. M. Ritchie

Cyanide owes its great success as a lixivant for gold to the great strength of the Au-CN bond. A replacement for cyanide must similarly be a strong complexant for gold (I) or gold (III). The various alternatives are briefly considered and it is concluded that only thiosulfate has a reasonable chance of replacing cyanide as a relatively non-toxic lixivant at the present time. However, there are considerable technical problems to be overcome before thiosulfate can be considered to be a realistic alternative to cyanide. Firstly, the leaching of gold using aerated ammoniacal thiosulfate with copper as a catalyst is a complex process which needs to be understood and controlled carefully in order to achieve maximum recoveries with minimum reagent losses. Secondly, there needs to be an option similar to the CIP process for recovering and concentrating the dissolved gold thiosulfate complex from pulps produced in the processing of low grade gold ores. The prospects for overcoming these technical' challenges are evaluated and discussed.

Jan 1, 1998

By Y. H. Zhan

A new P(MAAA) latex was developed and used in separation of ultrafine apatite from calcite and silica by means of selective flocculation. The latex was found to flocculate selectively apatite from apatite/calcite and apatite/silica mixtures in the presence of sodium tripolyphosphate and sodium silicate over a pH range from 9.1 to 9.4. The separation of apatite from calcite was more selective with sodium tripolyphosphate, while the results of the separation of apatite from silica were better when sodium silicate was used. In the apatite/calcite system, 90% apatite recoveries into a concentrate containing 34% P2O5 were possible in the presence of 4 mg/l of sodium tripolyphosphate at a latex dosage of 2.2 kg/t. In the apatite/silica system, 80% of apatite could be recovered into a concentrate containing 33% P2O5 in the presence of 20 mg/l of sodium silicate at a latex dosage of 1.1 kg/t. Deposition tests were used to study the interaction between the latex and the mineral surfaces. The results revealed that deposition density of the latex on apatite was higher than on calcite; the differences between the deposition onto these two minerals were higher at pH 9.2 than at 7.2. The results point to the stronger interaction between the latex and apatite surface when compared with the latex-calcite interactions.

Jan 1, 1998

By M. Ide, K. Takemura, S. Taenaka, K. Shinzawa

"The laser characteristic is strongry dependent upon the stoichometry of forsterite (Mg2Si04) and YSO (Y2Si05) single crystals, because of change in the valency and occupied sites of doped Cr in the lattice so that mole ratios of Mg0/Si02 and Y203/Si02 in the above single crystals must be precisely analyzed. In this study ICP-AES was applied for analyzing the mole ratios. Emission intensities corresponding to Si and Mg or Si and Y were measured at the same time with internal standard method. Preparation of sample solutions, measurement conditions of ICP-AES and interferences by coexisting components were discussed, resulting in excellently precise detennination of stoichometry of the above single crystals.IntroductionIn conventional stoichiometric analysis of single crystals, the stoichiometry is detennined by analyzing each component separately with titrimetry and gravimetry techniques. On the other hand, by stoichiometric analyses using ICP-AES, emission intensities corresponding to components are measured at the same time with internal standard method. Consequently erros caused by variation of analytical condition are canceled out.Cr-doped Mg2Si04 and Y2Si05 syngle crystals are known as laser materials. The change of valency and occupied sites of doped Cr affect the characteristic of single crystal as aleser materials. The change of state of doped Cr.Stoichometric composition of Mg2Si04 is constructed by 2mole of MgO and 1mole of Si02, and Y2Si05 is constructed by 1mole of Y203 and 1mole of Si02. In this study, required analytical precision in component mole ratio is within an error of 0.0010 for Mg2Si04 (CV=0.05wt%), and within an error of 0.0030 for Y2Si05 (CV=0.05wt%) (Table I )."

Jan 1, 1998

By B. T. Dawson

In today's work world, constant changes are a fact of life. Downsizing and technological advances are two of the most prominent concerns for employers and employees alike. In this article, the author addresses some of the emotional consequences of these changes and offers a new, empowering perspective for coping with them.

Jan 1, 1998

By Gus Van Weert

Solvent extraction of iron from zinc refinery liquors has been investigated before, mainly on the MEHPA / HC1 stripping combination. Leakage of chloride into the zinc electrolyte and the costs of handling the ferric chloride eluate have prevented this option from being adopted. This work explores the stripping of the ferric loaded organic phase with nitric acid to take advantage of the possibility of regenerating up to 6N nitric acid, and clean spheroidal hematite production by autoclave treatment of ferric nitrate solution. The work was carried out with DEHPA (di 2-ethyl-hexyl phosphoric acid) in kerosine. Degradation testing proved the DEHPA to be very stable, when exposed to nitric acid solutions up to 6N and temperatures up to 70" C over a period of five days. Ferric iron was removed completely froma synthetic hot acid leach liquor, irrespective of the DEHPA concentration. Stripping the ferric turned out to be the problematic step. Although the stripping efficiency was considerable better at 50 than at 20° C, it decreased with increasing DEHPA/Fe ration in the organic phase. Eluates of no more than 5-10 g/l ferric nitrate were produced. surprisingly, nitric acid was found to be a less efficient stripping agent than sulphuric at equivalent normality and its efficacy levelled off above 4N. It was concluded that DEHPA does not appear to be the S.X. reagent to take advantage of the low cost autoclave decomposition of ferric nitrate opportunity in zinc hydrometallurgy and that another S.X. reagent, also resistant to nitric acid, but based on weaker compound formation is needed.

Jan 1, 1998

By Tony van der Steen

Subsea mining system are designed for the following, principally different, tasks: Dreding - The destruction and removal of seabed sediments overlaying the bedrock. Cleaning - The competence of the system to systematically remove the deposits from an undulating hard bedrock with irregular pulleys, channels and fissures. Factors that influence the choice of dredging equipment for a specific project are project requirements such as environmental conditions, location and water depth. However, the most important selection criteria are soil conditions. As a consequence, possible dredging systems, must be evaluated for the following, principally different, characteristics: Dredging proficiency - The destruction of the strata by the equipment. System stability - The capability of the soil to support the subsea equipment The above parameters are determined by the strength characteristics and the bearing capacitv_ of the soil. The author describes the processes involved in excavating the subsea deposits and (transporting the gathered soil as a water-soil mixture to the surface using different excavation and transport methods.

Jan 1, 1998

By Stanton Johnson

Africa has a long history of mining and has been one’of the world’s greatest sources of gold and other mineral resources since the days of the Egyptian pharaohs. Mining in Africa has been dominated by South Africa since World War II. However, in the past 10 to 15 years, other African countries are beginning to slowly catch up with South Africa. Ghana, formerly called the “Gold Coast” is located on the West Coast of Africa. Ghana has been involved in mining activities for well over 100 years. Most of the mining activities centered on the mining of gold and were to a large extent limited to underground mining methods. With a more liberal economic policy regarding the exploration and exploitation of Ghana’s mineral resources, mining activities are beginning to evolve to a much larger scale. Drilling and blasting which is an important aspect of this evolution is also changing due to the involvement of foreign-based mining and explosives companies.

Jan 1, 1998

By L. Katgerman, H. J. Brinkman, J. Duszczyk

"The phase formation sequence in the in-situ synthesis ofTiB2 in A1 from elemental A1, Ti and B powders is identified with a combination of DSC and XRD. After the melting of A1, A1B2 and A13 Ti are formed at the contact areas between A1-B and A1-Ti. The influence of variations in starting composition on the reaction kinetics is explained in terms of there effect on the different processes in the formation sequence. Parameters in the starting composition that were investigated are; fraction ofTiB2 forming reactants, non stoichiometric Ti:2B ratio's, and addition of alloying element Cu.It was found that a liquid Al phase is necessary for the in-situ TiB2 forming reaction from the intermediate A1B2 and A13 Ti phases. If all Al is consumed by the formation of A1B2 and A13 Ti the TiB2 formation is halted until the decomposition of A1B2 and A13 Ti .Overstoichiometric amounts of B are needed for complete conversion of the Ti, present in A13Ti, into TiB2 This is explained by the fact that some of the stable A1B12 phase is formed after the decomposition of A1B2•The addition of Cu leads to an increasing reaction rate in the A1-Ti reaction. A possible explanation for this is that the A1-Cu liquid penetrates more easily between the Ti and its surface oxide than the unalloyed A1 liquid."

Jan 1, 1998