Incorporation of Geometallurgical Input into Gold Mining System Simulation to Control Cyanide Consumption

The Alhué deposit (Melipilla, Chile) is an example of a hydrothermal Au-Ag-Zn(-Pb) vein system hosted within the volcanic rocks of the Las Chilcas Formation. The dominant ore minerals observed are free electrum and native gold associated with silver sulfosalts, and with magnetite and base metal sulphides, including pyrite +/− sphalerite-galena-chalcopyrite. The alteration assemblage in the veins mainly consists of quartz epidote-chlorite-actinolite with lesser smectite, amphibole, and calcite-kaolinite-garnet. Mineralized veins also contain variable amounts of base metals, some of which (e.g., copper and iron) are considered harmful to the extraction of precious metals. Iron and especially copper minerals are known cyanide consumers; ore type classification schemes that do not consider the detrimental effects of such mineralogy or process elements can ultimately result in metal losses from ore feed restrictions, as well as spikes in cyanide consumption and higher operating costs. Mineralogical and geological variation can nonetheless be managed by applying alternating modes of operation as demonstrated in this paper; the decision to switch between modes is governed by current and forecasted stockpile levels feeding into the process. Simulations based on experiences at the Alhué deposit are provided that demonstrate the importance of standardized operational modes and their potential impact on cyanide consumption control.

Analysis of the current state of technologies in the field of recycling technogenic gold-containing raw materials

In this article, we review existing approaches to recycling technogenic raw materials (ore dumps, metallurgical production slag, mill tailings of ore-dressing plants, etc.), containing non-ferrous and noble metals, which are accumulated in almost non-ferrous metallurgy industries. An analysis of existing technologies for processing technogenic raw materials (pyrite cinders and flotation tailings of concentration plants), which include enrichment, pyro- and hydrometallurgical and combined ways of extracting valuable components, was conducted on the basis of a review of published sources. It was shown that enrichment (screening, desliming in a hydrocyclone, enrichment using a concentration table, magneticliquid separation, flotation), pyrometallurgical and combined ways for extracting noble metals from this type of raw materials are unprofitable. The most satisfactory results were obtained using hydrometallurgical methods to extract valuable components from technogenic raw materials. Various solvents, such as sodium cyanide, thiocarbamide, sodium thiosulphate and sodium sulphite were tested as leaching agents. Cyanation proved to be the most effective way to extract noble metals from technogenic raw materials; however, this process is characterised by a high consumption of sodium cyanide. Therefore, it is of importance to discover an approach to extracting valuable components from such problematic products in order to make their processing more cost-effective by reducing cyanide consumption while maintaining gold extraction. According to the obtained results, gold-containing raw materials are promising in terms of extraction of nonferrous and noble metals using hydrometallurgical technologies. Future research should identify rational methods for processing technogenic gold-containing raw materials in order to make the technology more profitable for extracting valuable components.

Assessment of the Possibility of Hydrometallurgical Processing of Low-Grade Ores in the Oxidation Zone of the Malmyzh Cu-Au Porphyry Deposit

The geochemical features of the ores present in the oxidation zone of the gold-copper Malmyzh mineral deposit are characterized. This deposit is located in the the Nanaian region of the Khabarovsk Territory in the Russian Federation, and a technological assessment of ores present was carried out. The initial gold content in the oxidized ore is 0.1–1.2 ppm, and the leachability of gold from different size fractions by hot cyanide revealed the fact that the ore is not refractory and characterized by increased cyanide consumption. The results demonstrated a high gold content in fine size fractions. Different leaching reagents were applied for oxidized ore testing. These reagents were used both for oxidative pretreatment and leaching process. Activation of solutions was carried out by electrolysis and UV radiation. Different reagent combinations were tested in laboratory with the highest gold recovery (94%) provided by the chloride system with peroxide-sulfuric pretreatment of the oxidized ore. The same reagent combination resulted in 91–94% gold recovery during the column leaching of crushed (−10 mm) ore.

Experimental study on thiosulfate leaching of gold from a high copper gold concentrate

The conventional cyanide leaching process is used to extract gold from a high copper gold concentrate. Because the copper associated minerals consume sodium cyanide in large quantities, the cost of the reagents is high and the economic benefit is not ideal. At the same time, a large number of cyanide tail slag are produced, which brings a series of environmental problems. In order to solve the environmental problems caused by excessive sodium cyanide consumption and cyanogen slag, the feasibility of leaching gold by thiosulfate in copper ammonia system was studied. The gold leaching rate of thiosulfate was increased to more than 90% by using the direct thiosulfate leaching process and pretreatment thiosulfate leaching process, which was close to the gold leaching index of sodium cyanide at the production site.

Sº and jarosite behavior during recovery of values from the direct leaching residue of sphalerite using cyanide and glycine

This paper reports the effect of the components of a direct leaching residue (jarosite and elemental sulfur), on the recovery of valuable metals such as gold and silver. Leaching media such as cyanide and mixtures of cyanide with glycine were used to recover the gold and silver from the residue; however, a low recovery of these metals was obtained. The above due to the negative effect of its components which cause problems in the extraction process such as encapsulation of silver (due to jarosite) and the formation of thiocyanate and re-precipitation of silver (due to sulfur). Various treatments prior to leaching were tested, finding that when the residue is desulfurized with perchlorethylene and subjected to an oxidizing alkaline hydrothermal treatment, the gold extraction increased from 39.73 to 88% and the silver extraction of 64.76 to 94.29%. Additionally, it was determined that when cyanide is assisted by glycine, the latter decreases the cyanide consumption by inhibition of the dissolution of iron and sulfur in cyanide.

Studies on heap leaching of gold with the addition of sodium acetate as an intensifying reagent

One of the cheapest methods of extracting gold is heap leaching. However, the recovery of gold by this method is relatively low, compared with cyanidation of the crushed material, so the search for the ways to intensify leaching and increase gold recovery is an urgent task. Investigations on heap leaching of gold from the gold-bearing ore of the Sari Gunay deposit were conducted using a promising reagent sodium acetate to intensify the heap leaching process. The results of assay-gravimetric, chemical, mineralogical and granulometric analyses of oxidized ore are presented. The average gold content in the ore was 2.90 g/t. According to the electron probe analysis, gold in the ore is present in the form of fine (micron) inclusions in minerals and ore rocks. Comparative studies on heap leaching of gold from the crushed ore with a grain size of –20 + 0 mm with the addition of sodium acetate and without that were carried out. The degree of gold recovery with sodium acetate at a flow rate of 0.5 kg/t was 58.74%, that without sodium acetate was 54.69%, i.e. the addition of the reagent provides an increase in recovery of more than 4%. Leaching with the addition of the reagent also reduces sodium cyanide consumption from 0.65 to 0.59 kg/t. The research results have shown that sodium acetate can be used to intensify the process of heap leaching of gold when the ore size is –20+0 mm.

Comparative Leaching Tests of Gold from Unroasted and Roasted Pyrite Using Microwave Radiation

Recently, the application of microwaves as a source of energy to the extractive industry has received particular attention from a large number of researchers. During the mid-1960s to 1988 most researches focused on the behavior of microwave radiant materials. These research activities continue to be important as long as the fundamental data on microwave interaction with materials is important for the successful application of technologies. This article presents the comparative cyanide tests of unprocessed and non-fired pyrite as well as the technical indicators obtained in these tests. After performing comparative cyanidation tests on unroasted and roasted pyrites at 400 and 700oC it was found that there was no significant reduction in lime and cyanide consumption for roasted pyrite at 400oC but for the roasted pyrite at 700oC the consumption of lime is reduced from 110 kg�t-1to 10 kg�t-1, cyanide is reduced from 2.8 kg�t-1 to 1.76 kg�t-1, and yields of leaching increase from 50 % for unroasted pyrite at 81 % for roasted pyrite.

PENINGKATAN RECOVERY EMAS DI CARBON IN LEACH PLANT PT. ANTAM TBK UNIT BISNIS PERTAMBANGAN EMAS PONGKOR, JAWA BARAT (STUDI KASUS KADAR BIJIH 3,5 – 4,5 GPT DI PLANT 1)

ABSTRAK Sejarah pengolahan mineral emas di Unit Bisnis Pertambangan Emas (UBPE) Pongkor PT Antam Tbk menunjukkan adanya tren menurunnya kadar bijih. Meskipun demikian, metallugist di UBPE Pongkor PT Antam Tbk secara berkelanjutan mencari berbagai penyebab dan cara untuk meningkatkan nilai recovery Plant walaupun kadar bijih yang cenderung menurun. Berdasarkan hasil pengolahan data terhadap Plant 1 dan Plant 2 di UBPE Pongkor, menunjukkan adanya potensi perbaikan di area Plant 1 UBPE Pongkor PT Antam Tbk. Makalah ini akan berfokus pada pengurangan kadar emas dalam sand tailing saat kadar bijih antara 3,5 sampai 4,5 gram per ton (gpt) di Plant 1 dalam rentang 2 bulan. Berdasarkan hasil analisa data yang diperoleh dengan menggunakan metode Anova, diperoleh hasil bahwa fraksi halus, kadar sianida, dan waktu tinggal slurry adalah hal – hal yang paling berpengaruh dalam proses pengolahan emas di UBPE Pongkor. Serangkaian perbaikan dan inovasi dilakukan untuk menyelesaikan masalah ini, yaitu dengan cara memodifikasi distribusi ukuran grinding ball, mengatur waktu mixing dan injeksi sianida dalam tangki, serta mengatur waktu dan memonitoring waktu tinggal slurry di dalam tangki leaching dan CIL.Berdasarkan hasil perbaikan dan inovasi yang dilakukan, terlihat bahwa nilai recovery plant meningkat dari sebelumnya 89,4% menjadi 92,7%. Hal ini diperoleh dari nilai perbandingan distribusi ukuran bola 1 : 2 untuk grinding ball 40-60 mm dan 60-80 mm. Sementara itu, penyesuaian terbaik kadar CN adalah 600 - 650 ppm ditambah dengan optimalisasi injeksi sianida menggunakan pompa dosing sehingga dapat mengurangi waktu mixing sianida yang 3x sehari menjadi 1x sehari dan menurunkan kadar CN dari 1,23 menjadi 1,13 kg NaCN/ton bijih. Berdasarkan hasil pengaturan mill feeder dan pompa slurry, waktu tinggal di tangki leaching dari 51 jam menjadi 60 jam. Kata Kunci : Emas, recovery, fraksi halus, sianida, waktu tinggal slurry ABSTRACT History of mineral processing at Gold Business unit at PT Antam Tbk shows a consistent trend of decreasing grade due to intensive mining operations. Therefore, The Metallurgist PT Antam Tbk Gold Business Unit is looking for roots caused and solutions to raise up plant recovery although the grade of gold declined. After deep data analyzing process at plant 1 & 2, it’s shown that plant 1 at Gold Business Unit PT Antam Tbk have a room for recovery improvement. This research focus on reducing sand tailing to improved gold recovery at feed grade between 3.5 and 4.5 ppm at plant 1. Deep analysis was applied to processing big data from the plant using ANOVA Methode, it’s shown that fine fraction, Cyanide and slurry residence time were the major factor of gold recovery. Series improvement were applied to solved this problem, such as modifying grinding ball particle size distribution, adjusting cyanide feeding on leaching thank based on feed grade, and monitoring slurry residence time. It was observed that the plant plant 1 recovery was increasing from 89,4% to 92.7%. The best grinding ball particle size distribution was 1 : 2 for 50 & 80 mm grinding ball. While the best applied cyanide concentration was between 600 - 650 ppm with some modification to improve process control of cyanide feeding. It’s shown that cyanide mixing process was decreased from 3 times a day to only one a day and also the cyanide consumption was reduced from 1.23 to 1.13 kg/ton ore. Some modification also carried out to increase slurry residence time. It’s proved that by applying the improvement it’s shown that the slurry residence time rose from 50 h to 61 h. Keywords : gold, recovery, fine fraction, cyanide, residence time.