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By Garry M. Johnston, Matthew L. Sutcliffe, Nicholas J. Welham

"The predominant current processes for the treatment of oxidized copper–cobalt ores use a reductive acid leach to bring the desired metals into solution. However, the indiscriminate nature of acid leaching results in a solution, akin to a metal soup, containing not only the Cu and Co but also an extensive suite of other metals. These other metals, notably Fe, Mn, Al, Ca, Mg and Si, all have to be eliminated from solution before the copper and cobalt can be recovered. Generally precipitation processes are used, but these tend to result in significant losses of copper and/or more usually cobalt from solution, leading to reduced recoveries. The AmmLeach® process uses ammonia-based chemistry to selectively leach the target metals, leaving the rest behind in the leach residue. This approach eliminates any need to clean the solution of unwanted metals and the consequent metal losses. After solid–liquid separation, the leach solution is immediately subjected to separation and concentration of the copper and cobalt using solvent extraction and electrowinning. INTRODUCTION The Copper Belt extends through the northern part of Zambia and the southern province of Katanga in the Democratic Republic of Congo (DRC) and contains the richest and highest grade deposits of copper and cobalt remaining in the world. Whilst there are a number of mines that produce sulphide concentrates, using froth flotation, the largest deposits are those where near-surface sulphides have been weathered over time to form oxides. The oxides cannot be concentrated by conventional physical means to give high recoveries. Despite this, a number of dense medium separation plants have been installed and used to produce a higher grade concentrate which is typically exported (while the rejects, typically 2 to 3% Cu are stockpiled for treatment at a later date). These plants generally have low recoveries and can only be used on the dwindling supply of high grade materials. In exporting the concentrate, little value adding is undertaken within the country, thus greatly reducing the economic benefits to the country. Recent changes to the regulations in the DRC will effectively ban the export of low value ores, concentrates and intermediates in favour of high value metals. This will necessitate mines to install new processing circuits in order to comply with these regulations. The head grade of ores currently being treated in the DRC is slowly decreasing as the extremely high grade materials are being selectively mined out. In many cases, the removal of a small volume of high grade materials leaves a much larger volume of economically marginal ore behind. This is due to a desire for short-term profits overtaking long-term prosperity. The remaining ore, which in many cases is still of relatively high value by world class standards, is likely to be only treatable by low cost heap leaching. There are some initial trials of heap leaching within the DRC, but implementation at a large scale has yet to be adopted."

Jan 1, 2015

By K. Wongaew, R. F. Cook, J. E. Carlson, T. Cleys, C. R. Smith

Steel fiber reinforced concrete segmental rings are the final tunnel lining for approximately 85% of the alignment for a 6.706-meter (22-foot) internal diameter tunnel being built as part of the East Side Combined Sewer Overflow Tunnel Project. This paper describes the approach adopted for the design of the steel fiber reinforced concrete segments, the development of the performance criteria for the segments, and the testing program used to confirm the performance of the materials and segments in the laboratory and in the field.

Jan 1, 2007

By Madgwick J. C

Leaching column techniques and microbiological methods have been described in detail in Madgwick and Ralph (1980) and Madgwick, Ralph and Sadler (1981).Sixty-nine kilograms of copper porphyry ore (0.4 per cent Cu, 2.7 per cent Fe) (particle size 2.4-11.2 mm; 60 The effect of sodium chloride (O.6M) irrigation on bacterial leaching of Bougainville copper porphyry ore was studied.Viability of suspended Thiobacillus ferrooxidants was reduced to less than 10 cells/ml in NaCI (O.6M) compared with 10' eells/ml in dilute II,SO, (0.01 M, pll 2). Salt irrigation had no significant effect on heterotrophic and sulphur oxidizing bacteria. Increascd short term recovery of copper (145 g/tonne/ 15 days) was followed by a marked decrease in tenors to less than 20 mg/litre/day compared with an acid leach, steady release, of 110 mg/litre/day. Resumption of acid leaching restored average solution copper to 46 mg/litre/day and the viability of suspended T Ierrooxidates to 10 cells/ml.Quantitative removal of sodium ehloride from ore particles was difficult and t races of NaCI remained for more than 200 days.

Jan 1, 1982

By Richard D. Begley

This paper summarizes the development of a mechanistic model to predict longwall mining induced surface subsidence. The model was originally developed from data acquired from a recent field study conducted in Northern West Virginia. Additional data has been collected from published sources and used to refine the model. The original model utilizes a negative exponential equation to predict the subsidence profile after the maximum possible subsidence is predicted based on the following input parameters: mine geometry, angle of draw and expansion factors for the deformation zones that have been identified above a longwall excavation. The model refinement involved the development of relationships between physical properties of the overburden strata and angle of draw in addition to the expansion or bulking factors. Therefore the refined model now requires only mine geometry and a typical geologic column from the site to predict the subsidence profile. This paper provides the methodology that enables the model to determine the angle of draw and expansion factors breed on physical properties of the site geology i. e. percent of strong rock and Rock Quality Designation of the overburden strata. In addition :comparisons of predicted results with measured data for over H) different sites with various geological and geometrical parameters located throughout the United States will be presented.

Jan 1, 1989

By Robert W. Bergwall

The $58,000,000.00 Columbia Heights Metro Station project (Section E3-a) forms a part of the Washington Metropolitan Area Transit Authority's (WMATA) Greenbelt Route - Green Line subway system. The Greenbelt route is one often segments that make up WMATA's subway system. Construction of the "cut and cover" reinforced concrete station includes 600 linear feet of an underground center platform, two ancillary service areas totaling 466 linear feet, two surface entrances, and vent shafts (See Figure No. 1). Station subgrade is approximately 60 feet below existing grade. The project can be divided into five main areas: 1. North Service Rooms and Passageway; 2. West Entrance and West Passageway; 3. East Entrance; 4. Center Platform; 5. South Service Rooms. Station dimensions are approximately 1068 feet long by 64 feet wide. Construction commenced at the site in January of 1995 and is expected to be completed by December of 1999. Franki Northwest Company, Crofton, Maryland, was awarded a subcontract by general contractor Blake Construction Company, Washington, D.C, to install the support of excavation and temporary decking systems. Franki's scope of work included the installation of 490 drilled soldier piles plus two elevator shafts totaling 29,700 linear feet of drilling; 1,150 tons of steel decking covered by 37,000. square feet of precast concrete deck planks; 120,000. square feet of timber lagging; 2,500. tons of struts and wales; and temporary support of underground utilities. In conjunction with Franki's work, Bencor Corporation of America, Dallas, Texas, constructed parallel, 600 foot long slurry walls which serve as permanent walls for the center platform portion of the station.

Jan 1, 1997

By C. Sun

The Leeville underground mine has been in operation since 2006 and currently produces 3,500 tons of ore per day with an average gold grade of 0.34 opt. The mine has a large horizontal extent of 2,500 feet by 500 feet and a vertical extent of 20 to 200 feet. Based on the characteristics of the ore bodies, a long-hole stope and cut-and-fill mining methods are used at the deposits which are divided into several mining zones. Cemented rock fill (CRF) is used as the backfill in the open stope. CRF design includes the aggregate design, the ratio of cement and fly ash design, and strengths for the long-hole stope and cut-and-fill. In the aggregate design, the particle size distribution is one of the most influential aggregate characteristics in determining how it will perform as a basic backfill material. Gradation helps determine durability, porosity, cement and water requirements, strength, and shrinkage. To cut the backfill cost, fly ash usually is added into the binder to replace some part of cement. However, adding fly ash can affect the strength of the backfill. The reasonable percentage of fly ash is also important for making a good and low cost backfill. Based on the different mining method, the different backfill with different strength is designed. A systematic quality control procedure for aggregate and backfill is followed. After four yeas application, CRF system in Leeville achieved very good results.

Jan 1, 2011

By Martin Banda, Gift C. Chisakuta, Moses Chabinga

KCM made a major modification to the Nchanga Tailings Leach Plant (TLP) changing the circuit from feeding post leach thickeners in parallel with leach discharge for solid ?liquid separation to a Counter Current Decantation (CCD) configuration utilising most of the existing equipment. The new circuit is aimed at improving copper recovery from approximately 70% to 79% through enhanced wash efficiency. The re-introduction of CCD circuit at elevated throughput of 46 000-tpd design was made possible due to the EIMCO E-Duc technology design which is an auto-dilution system that enhances solids settling rate. Commissioning (manual operation) which accommodated on-going operational requirements was phased and completed in January 2004. The circuit stabilised in subsequent months and in March alone an all time record copper recovery of 81.8% was achieved which contributed an extra 779 tonnes copper above the plant?s projected production plans. 100% of the project cost was paid back by end of April 2004 This paper discusses the work involved and the challenges faced during the re-commission

Jan 1, 2005

By V. J. Parks, Leonard Obert, A. J. Durelli

The state of stress in the region where core discing initiates has been investigated through the use of three-dimensional photoelastic models and the results of this study have been compared with those of a previous investigation of rock models. This comparison showed that discing initiates at a point of maximum shear stress, but that the magnitude of the shear stress, as determined photoelastically, is much larger than the shear strength of the rock as deter-mined from conventional triaxial testing. Moreover the shear stress required to produce discing is not constant, but depends on the applied stress field. Possible explanations for these effects are included in this article. In both exploration core drilling and overcore drilling to determine the stress in rock, occasional sections of the core obtained are broken into a series of discs.I* This discing phenomenon has been investigated in the laboratory to ascertain the state of stress required to produce it. The objective of the study was to obtain information that would make it possible to approximate the state of stress in subsurface rock from which disced cores are obtained. Referred to as the discing test, this laboratory study was conducted to determine the applied axial and radial loads required to produce discing in cylinders of rock from several sources. The dimensions of these cylinders are given in Fig. 1. Discing was produced by first applying an axial load to the cylinder and then diamond drilling the core along its axis. The radial load was then incrementally increased until discing initiated. The discs fractured or ruptured* on surfaces approximately normal to the axis of the core (planes A, B, C, D, E, Fig. 2). In many instances, the bottommost disc fractured before the drill had advanced past the points F and F', cre- ating a mush room-shaped fracture surface FEF'. Alternatively, it was also found that the same shaped fracture surface could be created by first drilling in an unloaded core to some point, say F and F', and then triaxially loading the core until fracture occurred on the plane FEF'. Jaeger and cook4 produced similar fracture surfaces in specimens subjected only to a radial biaxial load. From these laboratory studies and observations made in the field, it can be inferred that fracture or rupture initiates in the proximity of the lowest point in the kerf area, (at F and F', Fig. 2) and that the fracture or rupture surface progresses inward from this point. The experiments described above, however, do not provide any information as to the state of stress at the point at which fracture or rupture initiates. In the investigation discussed in the following sections, a three-dimensional photoelastic analysis was conducted to determine the stress distribution in the proximity of the kerf area of a triaxially loaded plastic (epoxy) cylinder containing a short section of core.

Jan 1, 1969

By Hookings H. J

UNDER a contract relating to the sale of a Broken Hill concentrate, the differentiation between granular and slime concentrate is determined by the proportion of the material which will pass a 200-mesh (I.M.M. standard) screen, and this has led to considerable attention being devoted to the question of laboratory sizing tests with screens of this order.It was found that many of the screens turned out by the makers during the war, and also since its termination, did not conform with the I.M.M. specifications as closely as some supplied in pre-war days. A perfect screen is hardly to be expected, but when the above contract was framed the parties anticipated that, by special care being taken during manufacture, a screen at least as good as the best pre-war screens would be available. This, however, has not proved to be the case, and the large irregularities in screens recently supplied render them useless for sizing determinations under the contract. The makers were approached through the Institution of Mining and Metallurgy, London, with the view of obtaining a few screens conforming as closely as possible to the 200-mesh I.M.M. standard, and in response to a request twelve pieces of screen were...

Jan 1, 1922

By P. B. Demarsh

"Location: - Mine village of Perron, 11 miles east of Val d’Or in northwestern Quebec.Life; - Started production 1932, ceased 1951.Ore: - Flat lying quartz tourmaline veins within granodiorite. Gold values related to coarse segregations of pyrite. Specific gravity 2. 80.Capacity: - 30 tons 1932-35; 125 tons 1935-37; 350 tons 1937-51.Assay: - Heads 0.21 to 0.34 oz. per ton. Grade obtained by mining 600 tons daily, sorting belt reject 250 tons, balance to milling; tailings loss 0.003 to 0.005 oz. per ton; extraction 9 8. 5%.Grind: - 70% minus 200 mesh. Detention time 26 hours.Treatment: - Cyanidation process.Reagents: - Per ton, aero cyanide (NaCn) 0.36; unslaked lump lime 1.50; zinc 0. 03; lead acetate 0. 006; steel consumption 2. 50.Water: - Equal mixture of swamp creek water and underground seepage.Cost: - Per ton, sorting and crushing 0.21; grinding 0.31; cyanidation 0.11; filtration 0.05; precipitation 0.02; miscellaneous 0.22; total 0.92.Power: - Hydro-electric, 25 cycle, 12, 000 volts reduced to 550 volts.Labour: - 95% Maritime and European, balance Quebec and Ontario.Tailings: - Absorbed over large swampy area on barren mining claims.Personnel: - 1 7 men including supervision.Flow sheet: - Canadian Mining Manual 1947"

Jan 1, 1966

By Troy Williams, Chris Preston, Ian Lipchak

"Underground blasting operations are challenging from the standpoint of the distribution of explosivesenergy representative of ring blasting. Energy from both shock and pressure regimes of commercialexplosives may appear concentrated in the collar region of a typical ring blast and diluted at the toes of holes due to the oblique geometries of blastholes. The non-homogenous nature of ore in whichexplosives are distributed via drillholes, adds to the complexities of generating particulate profiles from fragmented material with consistencies that are predictable from blast pattern to blast pattern - well suited for specific underground handling equipment and mill processing. In an ideal world, it would be the blasting operations themselves that represent the primary crushing mechanism, or at least mitigate mechanical crushing that can comprise a large component of the cost in generating suitable muck. This paper presents a dynamic break view in 3D that allows a planner to visualize the potential break zone around a blasthole generated by an explosive load using a Kleine field. Simple as it sounds, this methodology provides information that can be used in conjunction with cavity monitoring surveys (CMS) to potentially judge dilution due to overbreak as well as recovery for a typical blast. As examples, there are two break geometries that are examined regarding circular breaks and elliptical breaks around blastholes. Using a Kleine field to define break, a planner generated isosurface can be generated and compared to CMS data for calibration and prediction, using AEGIS 3D ring design software."

Jan 1, 2016

By McIntosh A. H, Torlach J. A

Since the last paper on mining methods was presented to the Annual Conference of the Australasian Institute of Mining and Metallurgy in 1:960, numerous advances have been made in drilling and blasting techniques at Mount Morgan. This paper describes the drilling and blasting methods being employed in the open cut. The change over to track mounted percussion drills for all types of primary drilling is almost complete and new techniques such as detergent drilling and pressure loading of ANFO are being applied successfully.

Jan 1, 1965

By J Fut=

The rotary drilling of rocks is widespread technology in mining and in tunnelling. The paper describes the results of research by aim to find relationships between optimal drilling and sound signal of drilling machine. The optimisation criterion has been formulated as the minimisation of specific energy with maximal effective speed of drilling in process of the rock disintegration. The research in this field has been made on experimental drilling stand. Our monitoring system has measured following variables: the pressing force of core drill, the revolutions of drilling head, the length of drilling, the time, the power, and the acoustic signals. For measurement of acoustic signals have been used very sensitively sound meter MEDIATOR 2238. From this measurements have been processed acoustic spectrums. Then we are found relationships between optimal control of drilling and composition of acoustic spectrum. On base of experimental results, we can formulate the optimal control following: æTo achieve the minimum difference between the equivalent sound level at the load and idle modes only at the representative frequency (f = 1000 Hz)Æ. The representative frequency dependences on type of the rock New method based on these relationships enables the optimal control of drilling machine by using measured acoustic signals only.

Jan 1, 2003

Fire suppression systems are now recognised by both NFPA 502 and PIARC as legitimate tools for both life protection and asset protection. The effectiveness of these tools for life protection demands extremely rapid and accurate incident detection coupled with rapid response. This rapid response must be integrated with other emergency systems such as emergency ventilation. The March 2007 fatal Burnley Tunnel incident in Australia graphically illustrates the potential of effectively integrated fixed suppression systems to protect people from harm and save infrastructure from damage. Despite the raft of technology available in that instance to the operator it was the human operatorÆs ability to effectively control the emergency equipment and not the equipment itself which so positively influenced the threat to life and property.

Jan 1, 2008

A methodology has been developed and demonstrated to be effective in reclaiming surface-mined land for a forestry post-mining land use. The Forestry Reclamation Approach (FRA) has been applied, almost exclusively, to flat or gently rolling surfaces. One of the primary steps in the FRA is to loosely grade the top 1.22-m (4-ft) of soil or substitute root growth medium to avoid excessive compaction. Concern has been expressed by regulators and the public that loose grading on steep slopes could lead to slope failures, which is one of the issues that SMCRA sought to correct. A field investigation was conducted at a steep-slope contour operation in eastern Kentucky to test the applicability of FRA in this setting. Final grading was performed according to FRA standards. The surface consisted of two distinct substitute rooting media. One was unweathered gray sandstone and the other was a combination of gray sandstone, brown sandstone, and topsoil. A variety of native hardwoods was planted following the FRA guidelines. The slope was instrumented with 70 survey monuments to monitor for mass movement. The reclaimed spoil was also characterized using bulk density and penetrometer resistance methods. The characteristics of both spoils are similar to those observed on flat to rolling surfaces where the FRA has been used and, after one year, approximately 70% of the trees had survived. There has been no major mass movement of the slope and slope stability analyses indicate that the application of the FRA has little impact on the overall stability of the slope.

Jan 1, 2011

By J Wallis

The stability of a raise bore is often assessed using the rock mass classification system published by McCracken and Stacey as a modification of the Q system. Often the assessment is done using a spreadsheet to manipulate geotechnical log(s) of boreholes. This paper discusses the inherent difficulties with this approach. Recent work by others to provide guidelines to the classification system is discussed. A recent example of the design and raise bore performance at Macraes Gold Mine, New Zealand is presented. A key issue is that a visual assessment of the drill core and/or core photographs is necessary to æground-truthÆ the classification.

Jan 1, 2008

By G. J. Ackland, K. Saksl, F. Brunke, M. Kohnke, B. Tegner, C. Siemers

The application of titanium alloys is limited to 550°C due to their poor oxidation resistance. It is known that the addition of niobium decelerates the oxidation of titanium alloys whereas elements like vanadium do not improve titanium’s oxidation resistance. The underlying mechanisms are not yet well understood. In the present study, different binary titanium-niobium and titanium-vanadium alloys as well as commercially pure titanium were investigated. Oxidation experiments were carried out at 800°C up to 288 hours followed by metallographic analyses to study the oxide layer morphology as well as the microstructure at the interface. Energy-dispersive X-ray spectroscopy mappings were used to identify possible changes in the distribution of the alloying elements. In addition, micro-focused hard X-ray experiments were performed to layer-wise analyse the phase composition in selected samples. As expected, the niobium containing alloys show a better oxidation performance compared to commercially pure titanium and the titanium-vanadium alloys. The oxide layer consisted of TiO2 and Ti2O3. Depending of the niobium content, a niobium-rich layer developed below the metal-oxide-interface minimising oxygen diffusion into the titanium matrix. Based on these results, the minimal niobium content needed for improved oxidation resistance was identified so that alloys of technical relevance can now be developed.

Jan 1, 2014

By G. W. Phelps

The field of industrial minerals is concerned with both sieve size and distribution of sub-sieve particles. A great deal of work has been reported on the techniques designed to provide information of this work. The term sub-sieve is here considered to include particular material that will pass a 325-mesh testing sieve. Plant men require rapid, precise and simple testing techniques that are, if possible, inexpensive with respect to equipment and manpower. A number of particle-size procedures have been developed, based on the Stokes equation for streamline movement of a particle of diam D, moving at velocity V through a liquid of viscosity n. 1 Particles are considered to be rigid, smooth spheres, completely dispersed at low concentration and moving at low velocity through the suspending medium with no slippage between particle and liquid. In particle size determinations the particle moves either from natural gravitational forces or those provided by centrifugal means. The diagram of Fig. 1 shows a spherical particle of density S and diam D moving downward at velocity L/T (or V) through a liquid of density So and viscosity n. The forces of gravity G (natural) or Gc (centrifugal) cause the particle to move through the liquid, thereby creating a viscous drag F or Fc. Within the streamline flow

Jan 1, 1967

"Mineral separation by exploiting the differences in mineral or metal specific gravities (sg) has been used since antiquity. In the modern era, this is most commonly seen in the recovery of gold by the use of centrifugal concentrators and the separation of coal (low sg) from waste rock (high sg) using heavy media cyclones and jigs. Other metals, notably oxides of tin and tungsten, are successfully recovered using some combination of jigs, spirals and shaking tables. However, many conventional gravity separation systems are limited to fairly coarse (>1 mm) to medium (>100 micron) particle separations due to the limitations of the gravitational separation forces in conventional equipment.Many of the high specific gravity minerals of interest are friable (e.g. cassiterite, wolframite, etc.) and excessive fines are generated in the grinding/comminution circuits which cannot be recovered with conventional equipment and may be lost directly to the tailings stream. Recent developments in centrifugal gravity separation equipment specifically designed for fine or ultra-fine particle separation and recovery have resulted in heavy mineral particle recoveries as fine as 10-20 microns. Ongoing research work has also identified the benefit of closed-circuit, multi-stage gravity recovery circuits using some combination of centrifugal concentrators and conventional equipment (tables, spirals, flotation, etc.) which can substantially increase overall plant recovery and final concentrate grade.This paper will provide details of the evolution of non-fluidized bed gravity concentrators for high sg mineral recoveries, correlate laboratory scale development work with plant results to confirm scale-up, and show how recent flowsheet developments have contributed to significantly improved plant performance.IntroductionGravity concentration is one of the oldest and most cost effective metal and/or mineral recovery systems, ranging from the classic hand panning of alluvial gold, through to the large scale production of a wide range of metals and minerals, including gold, silver, PGMs, black sands, coal and iron oxides. Conventional gravity recovery systems, comprised of some combination of sluices, spirals, jigs and shaking tables have limited application in fine particle (i.e. <100 microns) recovery, or where there are limited differences in particle specific gravities (sg), e.g. the separation of high sg, high value metal oxides (e.g. those of tin, tantalum or tungsten), from medium sg, lower value metal sulfides such as pyrite and pyrrhotite. These unit operations tend to have limited capacity, and while double and triple turn spirals and multi-deck tables have been developed, extensive floor space is often required in order to process the tonnages incorporated in the modern mineral processing plant."

Jan 1, 2017

By Sashir Estela De la Cruz

Actualmente la informática como herramienta aplicada al sector minero nacional, es importante -en tanto su empleo adecuado y racional implica alcanzar los objetivos planteados en forma más rápida y con mayor calidad. En este sentido, el presente texto describe el sistema informático que genera soporte al trabajo de modelamiento geológico para la estimación de reservas mineras, de forma que reduce el tiempo y personal empleado en dicha labor. Este sistema se divide en dos partes: estadístico y gráfico y determina la explotación de un prospecto de yacimiento minero.

May 25, 1998