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By E. Wolmarans

Dowding, Reynard and Associates Mineral Projects (DRAMP) was approached by the ARM/Lion Ore JV during July 2006 to conduct a feasibility study and control budget estimate for a new greenfield project for their NKOMATI Nickel mine 45 km east of Machadadorp. The scope included the evaluation and review of previous studies as well as a techno-economic trade-off on four possible comminution circuits ranging from conventional crushing and ball milling to fully autogenous milling. The following document encapsulates the thought process, decisions and the references used in selecting the final comminution route for the new 375 ktpm MMZ concentrator. Apart from the pilot test work performed, design simulations from a number of sources, performance results from a recently commissioned 100 ktpm MMZ concentrator, together with an understanding of the geology and mineralogy of the orebody and possible variations in the characteristics of the orebody over the life of mine were assessed. The results from all the above-mentioned work indicated that it would be possible to treat the MMZ ore at the desired throughputs to the desired grind sizes in a fully autogenous grinding (ABC-type) circuit. Autogenous milling is generally considered to be one of the higher risk milling circuits. This is mainly due to the fact that the ore, which is utilized as the grinding media, could be variable in terms of hardness and mineralogy/geology. Nevertheless, it is a very popular route as the cost benefits, more specifically the operating cost, are normally very attractive. This is especially the case when large low grade orebodies are evaluated. In North/South America, autogenous and semi-autogenous grinding circuits have been successfully employed during the past 20 years on large scale low grade operations. Taking all the above factors into account, the DRA/Nkomati project team recommended an autogenous circuit with the option to add steel to both the primary and secondary mills (if required) for the new operation in order to create the opportunity for significant operating cost savings. One has to bear in mind that all the new design simulations and decisions are based on previous work and potential unknown variations within the orebody will never be fully understood until the pit is operational for some time.

Jan 1, 2009

By C. E. Lesher

Two processes for agglomerating fine sized ores with low temperature coke are described. One process (Orcarb) agglomerates ores with limited amounts of carbon; the other (ore-carbon pellets) pelletizes fine sized ores, using low temperature cokes as the binder. Data are presented on the products obtained when taconite, magnetite, and hematite concentrates and several titanium oxide ores were used. REDUCTION of finely divided oxide ores has long been a major metallurgical problem. Various methods of agglomerating, notably briquetting, sintering, nodulizing, and pelletizing, have been developed and are in industrial use. Carbon is the usual reducing agent for oxide ores and attempts have been made to agglomerate fine ores with coking coal in order to get the metallurgical advantages of intimate contact and increased particle sizes suitable for subsequent smelting or reduction.1-3 For the past three years, research has been in progress on two processes for combining fine sized ores with reactive carbon in the form of low temperature coke. By one process, designated Orcarb, the ore is coated with coke carbon in amounts limited to that required for reduction and the result is an appreciable but limited increase in product particle size over that of the original ore. The second process, called ore-carbon pellets, produces pellets of fine ore bound by low temperature coke. The first objective was to develop a process for combining an oxide ore with only enough carbon for its reduction, the amount being that derived from the simple reduction equations. The carbon (that amount theoretically required for the reduction of the oxides in an agglomerate with ore), calculated to CO, ranges from 10 pct for zinc calcines or phosphate ores to 21 pct for rutile and 26 pct for silica. The carbon required for reducing iron oxides falls between the extremes—10 and 26 pct. When the carbon in the agglomerate is low temperature coke, reduction by hydrogen may be expected to take place, thus modifying the calculated carbon require- ment. On the other hand, in practice, an excess of reducing agent over that required in theory is normally used. Screen sizes of three iron ores on which the greater part of the experimental work was performed are given in Table I. Most of the testing was done with Freeport (Renton Mine) coal, but Pittsburgh and Elkhorn bed coals were also used. The first objective was to agglomerate the ore with sufficient low temperature coke to provide carbon for reduction, i.e., in the range from 10 to 25 pct. This was accomplished by preheating the ore, mixing with pulverized coking coal, and completing the coal carbonization in a plain steel rotating retort. The pilot retort, as it is now constituted, is shown by diagram in Fig. 1 and photograph, Fig. 2. Ore is fed into the upper kiln by a screw; it is heated to 900" to 1100°F by direct flame and it is then discharged into an insulating hopper from which it goes

Jan 1, 1956

By W. K. Cunningham, Edwin A. Gee

"INTRODUCTION This paper is one of many reporting on various aspects of the Bureau of Mines program during World War II directed toward alleviation of shortages in certain critical minerals some of these papers are published as technical papers information circulars, and reports of investigations, of the Bureau of Mines or in technical journals.The investigation described in this report was begun as part of the Bureau, of Mines program to develop acid processes for the extraction of alumina from low-grade bauxites and clays. A domestic shortage of high-grade bauxite suitable for treatment by the Bayer process became acute as a result of World War II; and intensive search for a practical acid process using lower-grade alumina-bearing raw materials was therefore warranted.A survey of the various methods for extraction of alumina from clay has been presented in an early Bureau of Mines publication. 4/ The principal barrier to the successful operation of an acid process has been the difficulty experienced in developing a suitable commercial method for purifying the crude aluminum salts after the clay extraction is complete. Iron salts are the foremost contamination influence. A review of the more prominent procedures that have been proposed for separating iron and aluminum salts can be found in the well-known text on aluminum by Edwards, Frary, and Jeffries. 5/The majority of acid processes to extract alumina from clay employ sulfuric acid as the leaching agent, and the purification is thus reduced to the separation of iron and aluminum, sulfates."

Feb 1, 1948

PRESENTATION PURPOSE ? Provide an overview of BC?s Environmental Assessment Actand environmental assessment process ? Identify key issues influencing environmental assessments in BC ENVIRONMENTAL ASSESSMENT ACT ? Major project proposals require an EA before a project can proceed ? Considers potential effects of proposed projects and ways to minimize or avoid adverse effects Minister of Environment and the Minister designated as responsible for project category decide whether or not to certify a project or order further assessment

May 1, 2006

I report results of over 850 microprobe analyses of sulfides, sulfosalts and electrum from the Favona adularia-sericite epithermal deposit of the Waihi district in the southern Hauraki Goldfield of New Zealand. Pyrite and marcasite from altered wall rocks that enclose the veins are stoichiometric, with compositions of FeS2, and no significant trace elements. In contrast, pyrite in veins commonly contains appreciable As (up to 3.2 wt per cent), and locally contains several tenths of a percent of Cu, Zn, Co, Ag and Sb. Vein pyrite may contain up to 0.22 per cent Au, but Au contents do not show a positive correlation with As. Limited electrum data indicate that it ranges from Ag0.47 Au0.53 to Ag0.54Au0.41Cu0.01Zn0.04. Chalcopyrite is stoichiometric CuFeS2, and most galena is PbS, but some galena contains up to 4.4 wt per cent Se, which substitutes for S in its lattice. The galena is not argentiferous; the highest Ag content is 0.45 wt per cent Ag. Sphalerite averages 4 mol per cent FeS, but ranges up to 15 mol per cent. Tetrahedrites [(Cu,Ag)10(Zn,Fe)2(As,Sb)4S13] show variable substitution, but most are Cu- Zn- and Sb-rich. Veins also contain aguilarite (Ag4SeS), stromeyerite (CuAgS), naumannite (Ag2Se), and mckinstryite [(Ag,Cu)2S].

Jan 1, 2004

By John E. Conley

The great demand for metallic aluminum for the prosecution of World War II caused the consumption of bauxite, the common ore of aluminum, to increase greatly. Prior to the war, the domestic deposits 'of bauxite, which occur chiefly in Arkansas and less extensively in Georgia, Alabama, and Mississippi, supplied only: about 30 ,percent of the ore required for aluminum-metal production and approximately 40 percent of alumina for all purposes. Enemy activities during the war seriously interfered with shipment of bauxite from the normal peacetime source in Surinam, Netherlands Guiana, South America, and focused attention on the necessity for increasing production from known domestic deposits. Even in peacetime, the production of aluminum ore in the United States could not satisfy requirements. Field drilling and exploration by Bureau of Mines engineers resulted in the discovery of important tonnages of high-grade bauxite suitable for treatment by the conventional Bayer process, Although an ore containing approximately 55 percent Al2O3 and only 2 to 4 percent SiO2 is preferable, it was necessary to use ores containing up to 8 or 10 percent SiO2. Because of the lower recoveries of aluminum with increased silica content, the Reconstruction Finance Corporation, through the Metals Reserve Company, established a base price of $4 for a 50-percent Al2O3 ore, containing 13-percent silica, with a premium or penalty of 14 cents for each percent of Al2O3 over or under 50; each percent of SiO2 under 13 was awarded a 20-cent premium, and each percent over 13 suffered a penalty of 43 cents; all prices were on the basis of 1 long ton of ore, dry basis.

Jan 1, 1949

By Nikhil Dhawan, Veeranjaneyulu Rayapudi

This study investigates upgrading of low-grade banded hematite quartzite iron ore (Fe ~31%). Conventional beneficiation was found to be futile. The susceptibility of iron phases to microwave exposure and their selective absorption assists in the liberation of iron values. Microwave exposure of coarse particles at 540 W for 10 min yielded a concentrate with Fe 56.30% and recovery of 50.68%. Conventional carbothermic reduction at 500 °C, 60 min and 9% charcoal yielded a concentrate with Fe 57.6% and iron recovery of 68%. The presence of sufficiently bonded silica leads to an easy formation of a fayalite phase. The microwave reduction design yielded FeG of 57.6%, FeR of 47% and a yield of 24% at an optimal condition of 540 W, 8 min and 6% charcoal. It was found that a small fraction of microwave-irradiated ore-charcoal mixture melted rapidly, and pure ferrite balls were observed within 8 min. An optical micrograph of a ferrite ball reveals the retained austenite and martensite phase.

By C. E. Lesher

Two processes for agglomerating fine sized ores with low temperature coke are described. One process (Orcarb) agglomerates ores with limited amounts of carbon; the other (ore-carbon pellets) pelletizes fine sized ores, using low temperature cokes as the binder. Data are presented on the products obtained when taconite, magnetite, and hematite concentrates and several titanium oxide ores were used. REDUCTION of finely divided oxide ores has long been a major metallurgical problem. Various methods of agglomerating, notably briquetting, sintering, nodulizing, and pelletizing, have been developed and are in industrial use. Carbon is the usual reducing agent for oxide ores and attempts have been made to agglomerate fine ores with coking coal in order to get the metallurgical advantages of intimate contact and increased particle sizes suitable for subsequent smelting or reduction.1-3 For the past three years, research has been in progress on two processes for combining fine sized ores with reactive carbon in the form of low temperature coke. By one process, designated Orcarb, the ore is coated with coke carbon in amounts limited to that required for reduction and the result is an appreciable but limited increase in product particle size over that of the original ore. The second process, called ore-carbon pellets, produces pellets of fine ore bound by low temperature coke. The first objective was to develop a process for combining an oxide ore with only enough carbon for its reduction, the amount being that derived from the simple reduction equations. The carbon (that amount theoretically required for the reduction of the oxides in an agglomerate with ore), calculated to CO, ranges from 10 pct for zinc calcines Or phosphate Ores to 21 pct for rutile and 26 pct for silica. The carbon required for reducing iron oxides falls between the extremes—10 and 26 pet' When the carbon in the agglomerate is low temperature coke, reduction by hydrogen may be expected to take place, thus modifying the calculated carbon require- ment. On the other hand, in practice, an excess of reducing agent over that required in theory is normally used. Screen sizes of three iron ores on which the greater part of the experimental work was performed are given in Table I. Most of the testing was done with Freeport (Renton Mine) coal, but Pittsburgh and Elkhorn bed coals were also used. The first objective was to agglomerate the ore with sufficient low temperature coke to provide carbon for reduction, i.e., in the range from 10 to 25 pct. This was accomplished by preheating the ore, mixing with pulverized coking coal, and completing the coal carbonization in a plain steel rotating retort. The pilot retort, as it is now constituted, is shown by diagram in Fig. 1 and photograph, Fig. 2. Ore is fed into the upper kiln by a screw; it is heated to 900° to 1100°F by direct flame and it is then discharged into an insulating hopper from which it goes

Jan 1, 1956

By K. B. Higbie

The quantitative separation of tantalum from columbium has been obtained through several liquid-liquid extraction processes. The valuable metals are separated by selective extraction from organic feed solution using a mineral acid or by selective extraction from aqueous feed solutions employing an organic solvent. The greatest degree of separation was obtained when organic ketones, particularly methyl isobutyl ketone, were used. Approximately 200 combinations of various organic solvents and mineral acids were tested as possible separation systems in which the tantalum or columbium content was extracted from organic solutions of the anhydrous pentachlorides.

Jan 1, 1956

By C. A. Rawlins

In an underground mine the ingress of heat can generally be attributed to two sources (Rock exposure and artificial mechanisms). The heat load relating rock exposure is subdivided into stoping [~40%] and haulages [~20%]. Rock is exposed to the ambient air and as a mine increases in depth the VRT (Virgin Rock Temperature) increases which leads to more heat ingress underground. To counter the heat ingress, especially in low-seam stoping areas of a deep level mine, air is, a) cooled artificially, b) the amount of rock exposed is reduced in the back areas of a stope by backfill applications, and c) ventilation air is used more efficiently. The heat reduction within a stope due to placed backfill could amount to ~50% for operations mining at a mean rock breaking depth of about 4500 mbc (metres below collar). Unfortunately, due to backfill transport methods, the physical fluid properties and application methods adopted, additional heat is induced into the air. The heat ingress due to backfill application could amount to as much as 25% of the overall in-stope heat load. This heat influx, seen as a negative aspect of backfill operations, is nevertheless overshadowed by the overall in-mine heat load reduction capability of about 20% attainable. The in-stope benefit [~50%] attainable from backfill operations could be increased with the incorporation of a heat exchange (cooling) methodology to offset the additional heat ingress into the surrounding air of a deep underground mine.

Jan 1, 2003

By R. Macleod, G. G. Bursey, T. R. C. Aston, J. E. Gale

In 1991 the problems of water inflow at the mine were addressed by a three-phase study. A three-dimensional finite-element computer code, CFEST (Coupled Fluid, Energy and Solute Transport), was used to model the system of groundwater flow around the mine, allowing assessment of the nature of the hydrological regime and of the various recharge mechanisms and sources of inflow. A regional (111 km2) model was used to simulate conditions of low to average mine inflow and provide boundary conditions for the subsequent local (16.8 km2) model. The regional model produced inflows of 0.26 m3/s and the local model 0.25 m3/s. With the use of constant head conditions to simulate a saturated flood plain along the Gays River and on doubling the rate of surface recharge, the local model produced a maximum mine inflow of 0.255 m3/s, in contrast to a measured peak mine inflow of 0.37 m3/s. The discrepancy suggests that during periods of high river level surface water gains access to features with high permeability, such as sinkholes, that may be connected directly or indirectly to the mine workings. Improved simulation of high inflow conditions can probably be achieved by determining the hydraulic conductivity of the river-bed and flood-plain sediments and modelling known sinkholes as discrete units. Possibilities for controlling the amount of inflow to the mine include sealing the flood plain with dykes, checking regularly for new sinkholes, which should be filled and capped, maintaining adequate mine drainage, and monitoring well levels and pressures on a continuous basis.

Jan 12, 1993

By J. Visman

The water cyclone washer has a number of advantages that makes it a desirable machine for the cleaning of fine-grained and friable materials such as 1,4 x 0" slack coals. In this paper results are presented of tests carried out on a shaly coal containing a high percentage of inter-grown materials from the Minto area, New Brunswick. The tests were con-ducted in the pilot plant of the Western Regional Laboratory in Edmonton. The test results show that the cleaning characteristics of this coal can be improved by lightly grinding the recirculating middlings fraction, which constitutes approximately 30 per cent (1.45-2.00 fraction) of the raw feed. • The cleaning of 1;4 x 0" fines by water cyclones compares favourably with other coal cleaning machines and has an added advantage of being extremely simple to operate, low in capital and operating cost, and high in capacity. The application of water cy-clones dates back approximately ten years, following its development at the Dutch State Mines. ( 3) A number of papers published d

Jan 1, 1961

By H. B. Hill

The Lake Creek gas-condensate field, Montgomery County, Tex., Was discovered early in 1941 by the Superior Oil Co. The field, shown in figure 1, is approximately 37 miles northwest of Houston, Tex., about 11 miles north of the Tomball, and 10 miles west of the Conroe oil fields. The field is on the downthrown side of a pronounced fault zone, and the general structure is anticlinal, with the long axis trending northeast and southwest. Production is obtained from multiple sandstone reservoirs in the Wilcox group of Eocene age between the approximate depths of 8,820 and 11,840 feet below the surface of the ground.

Jan 1, 1948

By Andrew Malinak, William Perkins

Seattle’s Elliott Bay Seawall protects the 1½ miles of downtown waterfront between South Washington and Broad streets. The seawall supports the urban waterfront and infrastructure, including the Alaskan Way Viaduct and Alaskan Way, access to historic waterfront piers, and numerous essential buried utilities. Most of the original seawall was constructed between 1916 and 1936 over soft/loose non-engineered fill, estuary, and beach deposits. The majority of the original seawall was a pile-supported timber relieving platform that extended up to 60 feet behind the face of the wall (Figure 1). With maximum spacing between the timber piles of 3 to 5 feet, a virtual forest of timber piles were driven for the original wall.

Jan 1, 2018

Mineral sand beneficiation plants comprise a complex combination of electrostatic, magnetic and gravity based separation techniques. These are combined and replicated in ways that provide the most efficient and cost effective separation of the valuable components. On line instrumentation has been developed at the JKMRC (eg. conductor/non-conductor(CNC) ) to measure the properties of the dry plant process streams. This paper provides results from the recent development of a novel low cost mass flow meter and discusses their combined use with the CNC instrumentation for the control of dry separation circuits.

Jan 1, 2003

By T Day, I G. Robertson, G J. Lyman

"Various approaches are used in the industry for establishing procedures (sample mass estimation and method of subsampling) for sampling and analysis of gold ores. The key factors that determine the accuracy of the results that can be expected for a given set of procedures are widely misunderstood.Recent developments in sampling theory show that the estimation of the accuracy that can be expected for a given gold ore as a function of sample mass is actually a simple matter. For a given sample mass, the accuracy that will be achieved depends only on the grade of the ore and the range of masses of the grains of the gold in the ore. This, of course, demands that a reasonable estimate of the distribution or spread of gold grain masses in the ore be known. But this is something that should be known if the gold is to be efficiently recovered in the processing plant.This paper provides a simple explanation of the real facts about sampling of gold, either in the field or in the plant, and provides results for a number of cases of combinations of fine and coarse gold in an ore. The prerequisites for accurate gold sampling are:correct sampling with no loss of sample or contaminationsample preparation procedures that recognise the impact of the gold grain mass spreadpractical analysis of the sample masses required to provide acceptable accuracy.Issues of corporate governance are also discussed.Citation:Lyman, G J, Robertson, I G and Day, T, 2016. The Simple Facts about Sampling Gold Ores, in Proceedings 13th AusIMM Mill Operators’ Conference 2016, pp 159–172 (The Australasian Institute of Mining and Metallurgy: Melbourne)."

Oct 10, 2016

By G. W. Weems

The Bureau of Mines used the Burnett method and apparatus to determine the compressibility factors for helium and carbon dioxide mixtures at 0°, 10°, 20°, 23°, and 35° C and to pressures of 900 psia. Tables of compressibility factors and approximations of the virial coefficients are presented for seven binary mixtures as well as for helium and carbon dioxide. The results compare favorably with existing data.

Jan 1, 1969

What is the BC AMTA? ?What is an ASEP? ?Who are the partners? ?Critical Success Factors ?Questions/Answers Aboriginal Skills and Employment Program (ASEP) ?A collaborative partnership of industry, employers, post secondary, First Nations and government ?Working together to plan and implement training and education programs ?Demand-driven: targeted JOBS for participants

May 1, 2010

By A Self, B Robertson

During 2015 and 2016, the authors conducted a number of surveys, workshops and reviews of ventilation and gas management practices in Australian underground coalmines. The study was funded by ACARP with a brief to identify gaps in ventilation and gas management practices for current and future needs, and to prepare a road map for change that would assist in delivering improvements.The study found that despite many examples of leading practice, gas and ventilation management is not afforded the attention and priority that is warranted and that this must be remedied if future operational performance targets are to be achieved and the safety of the industry protected. This paper summarises the findings of the study and presents the recommendations proposed.CITATION:Robertson, B and Self, A, 2017. A review of ventilation and gas management in underground mines, in Proceedings Australian Mine Vent Conference 2017, pp 277–284 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Aug 28, 2017