Comminution circuits are widely used in nonferrous and precious metals milling. Steel crushers, grinding mills and grinding media are typically used in comminution circuits. Comminution circuit wear products, in the form of metallic iron (steel) fines, are unavoidably produced and report to undersize products in the classification circuit, i.e., flotation or leach feed, while tramp grinding steel (TGS) reports to classification oversize products and returns to the grinding mill, building up in the circulating load. In laboratory studies, TGS reduced grinding efficiency significantly, the impact being dependent upon the size of the TGS present in the mill. Because iron fines adsorb flotation reagents effectively, iron fines increase reagent consumption and reduce flotation recovery. Iron fines are also capable of reducing gold recovery in cyanidation circuits via a cementation reaction. When using filtered gold solution from the cyanidation circuit of an Alaskan gold mine, the gold concentration in solution was reduced by 0.2 to 2.6% over various reaction periods when iron fines were present in the solution.

The free-milling ore characteristics of the Fort Knox gold deposit are ideal for studying gold and silver behaviors in various processing circuits without serious interference of sulfide minerals and other refractory factors. A galvanic effect leads to faster silver dissolution than gold dissolution and, as a result, the Au/Ag solution concentration ratio increases through the leach circuit. In the CIP circuit, Au adsorption is more effective than Ag adsorption. The Au/Ag solution concentration ratio decreases from 10.0 to 1.0 as the slurry advances from CIP tank #1 to CIP tank #5, while Au/Ag ratio in the carbon increases from 3.2 to 15.3, as the carbon advances from CIP tank #5 to CIP tank #1. Silver is stripped more effectively than gold from the loaded carbon. The Au/Ag ratio in solution reaches 2.0 in the first 75,700 L of strip solution that pass through the loaded carbon; as the stripping process continues, the Au/Ag ratio in the strip solution increases. In the electrowinning circuit, the Au/Ag ratio in the spent electrowinning solution is significantly lower than that in the electrowinning feed, indicating that Au is more rapidly deposited onto the cathode than Ag.

Deepwater discharges of sulfide flotation tailings have been practiced for several decades. Oxygen penetration and organic matter content in such environments are the two major factors that determine their reduction-oxidation condition and hence the rate of sulfide mineral dissolution under such discharge environments. In this study, the effects of solution redox potential and pH on the dissolution of some heavy metals from sulfide flotation tailings were examined. Zn, Pb and Cu concentrations in solution were all higher at a higher redox potential and lower pH. Dissolved Zn concentrations reached 100 ppm even at a low redox potential of 105 mV and near neutral pH after 33 days. Pb concentrations were below 2 ppm at 105 mV and near neutral pH after 165 days. At 105 mV, the Cu concentrations were below 0.2 ppm at near neutral pH after 165 days. At pH 2.3, the Fe concentrations were higher at 420 mV than at 105 mV. But at pH 6.8, the Fe concentrations were lower at 420 mV than at 105 mV, due to the oxidation of Fe2+ to Fe3+ and the subsequent precipitation of ferric hydroxide. Arsenic concentrations in solution were all less than 0.002 ppm. For Zn, Cu and Fe, no significant differences were observed between the tests in solutions prepared using seawater and those using distilled water. However, Pb concentrations reached higher levels in solutions prepared using seawater.

Residual cyanide in cyanidation tailings can be detoxified by its oxidation to cyanate, as in the Inco and Degussa processes. However, cyanate, the product of detoxification and other oxidation processes, may be reduced to cyanide in reducing environments, e.g., deepwater discharge of cyanidation tailings. This study shows that the reduction of cyanate to cyanide occurs and that the reduction rate increases with decreasing solution redox potential. Solutions with lower pH favor the reduction rate. No significant difference was observed between the results of tests prepared with distilled water and seawater, and the presence of cyanidation solid tailings does not show an effect on the reduction. Because cyanate can be reduced to toxic cyanide, causing environmental concerns, the deepwater discharges of cyanidation tailings must be handled carefully.

The effect of metallic iron fines on froth flotation of a sulfide ore was examined by comparing the flotation results with and without iron fines present in the flotation feed. Mild steel and ceramic laboratory ball mills were used to grind 600 g sulfide ore charges. The mild steel ball mill generated approximately 4.6 g of iron fines (0.8% of the ore charge mass) when grinding to a product size of 80% passing 200 mesh (75 /μn). When a low dosage of a weak xanthate collector was used in flotation, the impact of the iron fines on sulfide flotation was marked. Higher dosages of stronger xanthate collectors rendered the impact of the iron fines much less significant. Consumption of the collectors by the powdery iron fines is believed to cause the differences observed in the flotation results. A galvanic effect on flotation was not evident in this study. The impact of the iron fines on silicates depression was also examined.

The Fort Knox Mine is located in Alaska’s interior, where the average ambient air temperatures range from −24°C (−11°F) in January to 16°C (61°F) in July. The mill processes a free-milling gold ore utilizing both a gravity recovery circuit and conventional cyanide leach/carbon-in-pulp circuits. Mathematical models have been developed to accurately predict the impact of leach circuit slurry temperature on gold leaching, carbon adsorption and cyanide destruction kinetics. Additionally, an energy-balance approach has been used to model the seasonal variations in slurry temperatures throughout the Fort Knox mill. The energy-balance model, combined with the kinetics models previously developed to predict plant performance, was used to justify a tailing thickener installation at the Fort Knox Mine. This paper describes the development of the energy-balance model and its prediction of plant performance assuming a tailings wash thickener to reduce heat loss from the mill. Also presented is an analysis of the post-project plant performance that validates the model and shows that the mine produced an additional 596 kg (19,155 oz) of gold while reducing mill reagent costs by $6,302,000 during the first 38 months of operation.

The Fort Knox Mine is located in Alaska’s interior where the average ambient air temperatures range from −24°C (−11°F) in January to 16°C (61°F) in July. The mill processes a free-milling gold ore utilizing a gravity recovery circuit and conventional cyanide leach/carbon-in-pulp circuits. Plant slurry temperature cycles seasonally and allows for a unique opportunity to measure the impact that temperature has on gold leach kinetics, carbon adsorption efficiency and cyanide destruction reactions. Mathematical models were developed to accurately predict actual mill performance. This paper describes the development of the models and presents data, which allow for a better understanding of the impact of circuit operating temperature on process efficiencies.

The purpose of this study was to develop a mathematical model to describe the oxidation of cyanide with SO 2 , as proposed in the Inco Process. This research employed a direct method for measuring the change in cyanide concentration, with respect to time, as affected by varying pH, temperature and concentrations of SO 2 and copper. This model may have applications for determining optimum conditions for a process well known and used in the mining industry. It was determined that the reaction is first-order with respect to [CN − ], [SO 2 ] and [Cu +2 ]. The reaction order is 0.1, or zeroth order, with respect to [H + ], for a pH range of 8.0 to 9.5. The activation energy, E a , is 28.5 kJ/mol. Given D = ([SO 2 ] 0 − [CN − ] 0 ), the overall rate equation was determined to be ln{([CN − ] 0 ([CN − ] t + D))/([CN − ] t ([CN − ] 0 ) + D))} = ke −28500/RT D[Cu +2 ][H + ] 0.1 t, where the subscripts denote concentrations with respect to time (t).

A qualitative test program to study placer gold nugget recovery using an eddy-current separator was sponsored jointly by the University of Alaska and Eriez Magnetics. Various gold nuggets from Alaska, ranging in size from 6.4 to 63.5 mm, were tested. All tests were performed using an Eriez Laboratory ESC, Model REA, fitted with a 2,000-gauss, surface field-strength magnetic rotor. The belt speed was set to 91 ml min (300fpm) and the rotor speed was set to 2,500 rpm in the forward direction. The splitter was adjusted to just reject the largest pieces of rock, which averaged 76 × 102 × 127 mm. Based on the results of this limited test work, it is being recommended to placer gold mine operators that ECS technology is a viable and practical concentration technology for coarse gold recovery. Silver and platinum were also tested.

Fairbanks Gold Mining Inc. (FGMI) began blending and processing True North ore at its Fort Knox mill on April 12, 2001. Gold extraction subsequently dropped from 88% to 71% during the last week of April 2001. Twenty-five percent of the gold values remaining in the leach tailings were gravity-recoverable free gold. Significant quantities of stibnite and arsenopyrite were also identified in the tailings. In laboratory bottle-roll tests, the gold extraction from a composite of leach circuit feed collected during the same period was 72.6% when leached at an average NaCN concentration ofO. 16g/kg. The extraction increased to 82.4% with 0.5 g/kg NaCN and increased to 91.7% with 0.5 g/kg NaCN plus 0.075 g/kg Pb(NO 3 ) 2 . A passivating layer, resulting from sulfide mineral dissolution, surrounded the liberated gold particles and likely contributed to the reduced gold recovery. The gold recovery can be substantially increased by the addition of lead nitrate. Lead species remain in the solid residues after leaching and were not mobilized using meteoric water mobility procedure.

At the Fort Knox Mine, Fairbanks, Alaska, carbon fines are generated during stripping, acid washing and kiln regeneration operations. Because of their gold value, carbon fines are collected by a filter press. Samples of the carbon fines show that they consist of 20% to 30% carbon, 70% to 80% grit and 0.17 to 0.68 g/kg (5.0–19.9 oz/st) gold. About 80% of the gold values in the carbon fines used in this study are in the form of recrystallized free gold grains, with the balance being the adsorbed gold on the carbon. High recoveries of gold and carbon were obtained from the fines in laboratory tests using froth flotation, with a high rejection of the grit. More than 95% of gold and 90% of carbon can be recovered by three stages of gold flotation followed by two stages of carbon flotation.

A soil washing and leaching process was tested for removing lead from soils. A soil-washing circuit, including size and gravity separations, was employed to remove the coarse metallic lead particles, while the leaching was applied to remove fine metallic lead particles and other lead species. The soil-washing tests proved that the metallic lead particles larger than 0.15 mm (100 mesh) could be effectively removed. The sodium-chloride-based leaching solution with ferric chloride or sodium hypochlorite as oxidants was adopted in the leaching. The leaching experimental results indicated that under the pH of 2 and Eh of 1,300 mV, the metallic lead particles smaller than 0.15 mm and other lead species can be dissolved in the leaching solution within 60 minutes.

Over a period of eight years (1989 through 1997) the Mineral Industry Research Laboratory (MIRL) at the University of Alaska, Fairbanks, AK, has investigated the effects of temperature, residence time and particle size on the degree of low-rank coal (LRC) upgrading for a nonevaporative, hydrothermal drying process, also known as hot water drying (HWD). The following three Alaskan LRC s were used in this study: Usibelli coal (Seam No. 4) from the Nenana Coal Field, Little Tonzona coal from near McGrath, and Beluga coal from the Beluga-Yentna field on the west side of Cook Inlet. Replicated factorial tests were conducted. Experimental results indicate that most of the low-rank coal upgrading via HWD occurs rapidly within the first 10 to 20 min of residence time. Thereafter, the upgrading process slows down, and more than 100 min of additional residence time is required to produce percentage increases equivalent to what was achieved in the first 10 to 20 min. The properties of the hot water dried products were very sensitive to process temperature. Calorific value, carbon and oxygen contents, equilibrium moisture levels and Hardgrove gindabilities all appear to change linearly with temperature within the 275 °C to 325°C (525°F to 615°F) range.

This paper reviews the recent research on sluiceboxes, and based on this, presents a unified analysis of their behaviour. The available laboratory data. suggests that substantially lower d(A,50)'s could be obtained in commercial units. To obtain these low, competitive d d(A,50)'s, throughput must be reduced, and feed top sizes reduced Given that most placer deposits have only a limited range of gold particle sizes, these conditions are realistically obtainable.A comparison of laboratory and plant data on gold sluiceboxes suggests that to perform successfully, it must achieve four processes:i) deliver gold particles to the plane of the rifflesii) pull gold particles into the rifflesiii) hold gold particles in the riffles, in preference to other mineralsiv) transport gangue minerals along and off the sluice.Any one of these processes can be performance limiting. Under carefully controlled conditions, d(A,50)'s of the order of 50 mu m wry be possible - a performance similar to that obtained on a spiral.It is considered desirable that operators make an effort to determine the size distribution of the gold in their operations, and a method for determining this is proposed.

This project is investigating the suitability of hydrothermally dried low-rank coals for pulverized fuel injection into blast furnaces in order to reduce coke consumption. Coal samples from the Beluga coal field and Usibelli Coal Mine, Alaska, are being used for the study. Crushed coal samples were hydrothermally treated at three temperatures, 275, 300 and 325{degrees}C, for residence times ranging from 10 to 120 minutes. Products have been characterized to determine their suitability for pulverized coal injection. Characterization includes proximate and ultimate analyses, vitrinite reflectance, TGA reactivity and thermochemical modeling. A literature survey has been conducted.

This project is investigating the suitability of hydrothermally dried low-rank coals for pulverized fuel injection into blast furnaces in order to reduce coke consumption. Coal samples from the Beluga coal field and the Usibelli Coal Mine, Alaska, are being used for the study. Crushed coal samples were hydrothermally treated at three temperatures, 275, 300 and 325{degrees}C, for residence times ranging from 10 to 120 minutes. Products are being characterized to determine their suitability for pulverized coal injection. Characterization includes proximate and ultimate analyses, vitrinite reflectance and TGA reactivity. A literature survey is being conducted.

This project investigated the suitability of hydrothermally dried low-rank coals for pulverized fuel injection into blast furnaces in order to reduce coke consumption. Coal samples from the Beluga coalfield and the Usibelli Coal Mine, Alaska, were used for the study. Crushed coal samples were hydrothermally treated at three temperatures, 275, 300 and 325{degrees}C, for residence times of 10, 60 and 120 minutes. Products were characterized to determine their suitability for pulverized coal injection. Characterization included proximate and ultimate analyses, vitrinite reflectance and TGA reactivity. A literature survey was also conducted.

This project is investigating the suitability of hydrothermally dried low-rank coals for pulverized fuel injection into blast furnaces in order to reduce coke consumption. Coal samples from the Beluga coal field and the Usibelli Coal Mine, Alaska, are being used for the study. Crushed coal samples will be hydrothermally treated at three temperatures, 275, 300 and 325 C, for residence times ranging from 1 to 120 minutes. Products will be characterized to determine their suitability for pulverized coal injection. Characterization will include proximate and ultimate analyses, density measurements, reflectance and TGA reactivity. A literature survey is being conducted.

A column-flotation system was designed to conduct release analyses on a very finely disseminated, stratiform copper ore from the Denali Copper Prospect in Alaska. Chalcopyrite is finely disseminated in a carbonate matrix, and pyrite is disseminated in both the chalcopyrite and the gangue. Grinding to −20 µm is required for adequate liberation.